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IC 


9059 



Bureau of Mines Information Circular/1986 




Precious Metals Recovery From Low- Grade 
Resources 

Proceedings: Bureau of Mines Open Industry Briefing Session 
at the National Western Mining Conference, Denver, CO, 
February 12, 1986 



Compiled by Staff, Bureau of Mines 



UNITED STATES DEPARTMENT OF THE INTERIOR 



Information Circular 9059 

Precious Metals Recovery From Low- Grade 
Resources 

Proceedings: Bureau of Mines Open Industry Briefing Session 
at the National Western Mining Conference, Denver, CO, 
February 12, 1986 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 
Donald Paul Model, Secretary 

BUREAU OF MINES 
Robert C. Norton, Director 






Library of Congress Cataloging in Publication Data: 



Precious metals recovery from low-grade resources. 

(Bureau of Mines information circular ; 9059) 

Includes Bibliographies. 

Supt. of Docs, no.: I 28.27:9059. 

1. Precious metals— Metallurgy— Congresses. I. United States. 
Bureau of Mines. II. National Western Mining Conference ( 1986 : Den- 
ver, Colo.). III. Series: Information circular (United States, Bureau 
of Mines) ; 9059. 



<SN295.U4 [TN7591 622s [669'.2l 85-600302 



fe 



o 



PREFACE 



i 



Q- 



Most of the papers included in this Information Circular were present- 
ed at a Bureau of Mines open industry briefing held in conjunction with 
the National Western Mining Conference on February 12, 1986, in Denver, 
^ CO. The Bureau often sponsors meetings of this nature in an effort to 

move new technology into industry practice by drawing attention to de- 
velopments that may solve certain problems or improve upon current tech- 
niques. Those desiring more information about Bureau research programs 
should contact the Bureau of Mines, Branch of Technology Transfer, 2401 
E Street, NW, Washington, DC 20241. 






3 

3 



Ill 



CONTENTS 

Page 

Preface 1 

Abs tract 1 

Introduction. 1 

Ion-Exchange Research in Precious Metals Recovery, by Glenn R. Palmer 2 

Staged Heap Leaching-Direct Electrowinning, by C. H. Elges and 

M. D. Wroblewski 10 

Mercury Precipitation During Cyanide Leaching of Gold Ores, 

by Richard G. Sandberg 19 

Carbonaceous Gold Ores, by B. J. Scheiner 26 

Carbon Adsorption-Desorption, by J, A, Eisele 34 

Heap Leaching, by J. A. Eisele 37 

The Carbon-in-Pulp Process, By Stephen D. Hill 40 

Precious Metals Recovery From Electronic Scrap and Solder Used in Electronics 

Manufacture, by B. W. Dunning, Jr 44 





UNIT OF MEASURE ABBREVIATIONS USED IN 


THIS REPORT 


A 


ampere 


L/min 


liter per minute 


A/ft2 


ampere per square foot 


mg 


milligram 


A/in2 


ampere per square inch 


mg/L 


milligram per liter 


°C 


degree Celsius 


mg/min 


milligram per minute 


cm 


centimeter 


mg/(min«g) milligram per minute 








per gram 


cm^ 


cubic centimeter 










min 


minute 


ft 


foot 










mL 


milliliter 


g 


gram 










ml./min 


milliliter per minute 


gal 


gallon 










mm 


millimeter 


g/cm3 


gram per cubic centimeter 










mt 


metric ton 


g/L 


gram per liter 










pet 


percent 


g/mt 


gram per metric ton 










ppm 


part per million 


h 


hour 










st 


short ton 


in 


inch 










st/d 


short ton per day 


kg 


kilogram 










st/wk 


short ton per week 


kg/mt 


kilogram per metric ton 










tr oz 


troy ounce 


kW'h/st 


kilowatt hour per short 








ton 


tr oz/st 


troy ounce per short 
ton 


L 


liter 










V 


volt 


lb 


pound 










vol 


volume 


lbf/in2 


pound (force) per square 








inch 


wt pet 


weight percent 


Ib/st 


pound per short ton 


yr 


year 


lb/(sf h) 


pound per short ton 
per hour 







PRECIOUS METALS RECOVERY FROM LOW-GRADE RESOURCES 

Proceedings: Bureau of Mines Open Industry Briefing Session 

at the National Western Mining Conference, 

Denver, CO, February 12, 1986 

Compiled by Staff, Bureau of Mines 



ABSTRACT 

This information circular describes research the Bureau of Mines has 
conducted to improve technology for recovering precious metals from low- 
grade resources. Many of the reported findings are recent, and some of 
the work described is ongoing. Topics discussed include cyanidation of 
carbonaceous gold ores to enhance gold recovery, a new method for pre- 
cipitating mercury during cyanide leaching of gold ores, a staged heap 
leaching process to generate suitable solutions for direct electrowin- 
ning of gold, the use of anion-exchange resins to recover gold from cya- 
nide solutions, and precious metals recovery from electronic scrap. 

INTRODUCTION 

Although U.S. gold and silver resources are extensive, many of these 
resources are extremely low in grade and cannot be developed economical- 
ly with conventional mining and mineral processing methods. By devel- 
oping more efficient and economical mineral recovery techniques, the 
Bureau of Mines has helped to make these low-grade resources more acces- 
sible. Past research by the Bureau focused on the carbon adsorption- 
desorption, carbon-in-pulp , oxidation pretreatment , and heap leaching 
technologies used in various phases of precious metals recovery. Cur- 
rent research highlights even more economical and effective approaches 
to gold and silver recovery, such as the use of ion-exchange resins, 
staged heap leaching followed by direct electrowinning, and mercury pre- 
cipitation from leach solutions. The results of the Bureau's research 
in these areas, including both past and current work, are presented in 
the following papers. 



ION-EXCHANGE RESEARCH IN PRECIOUS METALS RECOVERY 
By Glenn R. Palmerl 



ABSTRACT 



The Bureau of Mines investigated the 
use of ion-exchange technology in the re- 
covery of precious metals from cyanide 
leach solutions. Strong-base resins such 
as Amberlite IRA-430 are nonselective, 
and using them results in high metal 
loading of precious metals and mercury, 
but generally they are difficult to 
strip. Research was conducted to devise 
a sequential stripping technique for var- 
ious strong-base ani on-exchange resins. 
The sequential stripping of IRA-430, for 
example, eluted 100 pet of the mercury 
with 2N H2SO4, 100 pet of the silver with 
200 g/L NaCl in IN HCl , and all 
remaining gold with 0.75 pc 
150 g/L plus 5 g/L NaOH. 

Weak -base resins generally are selec- 
tive and yield low metal loading, but are 



of the 
NaClO in 



not difficult to strip. Research was 
conducted to examine the possible appli- 
cation of various weak-base anion- 
exchange resins in the extraction of gold 
and/or mercury from a cyanide solution. 
Batch equilibrium experiments were per- 
formed at different pH values to evaluate 
which resins might show promising results 
in an actual mill circuit. Three experi- 
mental resins may have possible applica- 
tion because of their effectiveness in 
the pH range of 10 to 12. One of these 
resins, when tested with actual mill 
solutions in a flow-through column ex- 
periment, loaded to about 140 tr oz/st Au 
and produced a peak eluant concentration 
of 35 ppm Au. 



INTRODUCTION 



Numerous cyanide leach carbon-in- 
pulp (CIP) , carbon-in-column (CIC), and 
carbon-in-leach (CIL) milling facilities 
have been constructed to process low- 
grade gold-silver ore deposits throughout 
the Western United States (J_-9).2 The 
use of activated carbon has greatly im- 
proved the efficiency of gold recovery 
operations. However, several drawbacks 
are inherent with the use of activated 
carbon: 

1. Used carbon must be thermally re- 
activated to maintain its effectiveness 

M-ID • 

^Metallurgist, Salt Lake City Research 
Center, Bureau of Mines, Salt Lake City, 
UT. 

^Underlined numbers in parentheses re- 
fer to items in the list of references 
at the end of this paper. 



inhibited by 
organic re- 



2. Carbon loading may be 
the adsorption of CaC03 , 
agents, and clay-type minerals. 

3. Carbon stripping must be performed 
in a pressurized vessel at elevated tem- 
peratures, 120° to 130° C, to shorten the 
stripping time. 

An alternative recovery technology 
studied by the Bureau of Mines involves 
the application of anion-exchange resins 
for the recovery of gold from cyanide 
solutions. As early as 1949 (11-13) , 
attempts were made to use both weak- and 
strong-base anion-exchange resins. Since 
that initial work, various applications 
and problems associated with anion- 
exchange resins have been examined. 

Strong-base resins are generally less 
selective for precious metals, but 
have higher loading capacities and are 
much less affected by pH than weak-base 



resins, A general mechanism for loading 
a strong-base resin is as follows: 



I-NR3+X- + M(CN)2- 

^ |-NR3+M(CN)2" + X- 



(A) 



where |— , NR3 , H+X" , and M(CN)2" repre- 
sent the polymeric resin matrix, func- 
tional group, acid, and metal-cyanide 
complex, respectively. Weak-base resins 
are generally more selective for precious 
metals, but have lower loading capaci- 
ties. Because protonation is required to 
extract anions, a pH of less than 10 is 
usually required (14). A general mechan- 
ism for loading a weak-base resin may be 
represented by the following: 



I-NR2 + H+ + X^ 
^=> [-NR2H+X" 



(Protonation) (B) 



I-NR2H+X- + M(CN)2' 

^=^ |_NR2H+M(CN)2" + X- (Loading) (C) 

Both resin types can be eluted at ambient 
temperature and pressure. Weak-base res- 
ins generally can be eluted using a di- 
lute caustic solution. Strong-base res- 
ins, however, are more difficult to elute 
and require a more rigorous treatment. 
Potassium thiosulfate, acetone plus HCl, 
ethyl acetate plus HNO3 diluted with 
water (13) , zinc cyanide, and dimethyl 
formamide have each been used to recov- 
er gold ( 15 ) . Each elution technique 
has one or more of the following dis- 
advantages: fire hazard, noxious gas 
formation, or the need for special regen- 
eration procedures. The mechanisms for 
eluting strong- and weak-base resins are 
described by the reverse of reactions A 
and C, respectively. 



STRONG-BASE RESINS 



During the past few years, the Bureau 
of Mines has investigated problems as- 
sociated with the elution of gold, sil- 
ver, and mercury from strong-base resins 
such as Amberlite IRA-430 and 900 and 
Dowex 21-K, SBR, and SMA-1.3 Tests were 
initiated by loading resins with 145 tr 
oz/st each of gold, silver, and mer- 
cury from cyanide solutions doped with 
radioactive tracers for analysis. The 
loaded resins were eluted with several 
eluants, such as H2SO4, HNO3 , acid chlo- 
ride, hypochlorite, and/or alkaline chlo- 
ride solutions. 

Figure 1 shows a simplified flowsheet 
for the preferred eluting sequence. Ta- 
ble 1 presents the results of using the 
simplified flowsheet with each of the 
five resins. Generally, all five strong- 
base resins behaved similarly. Within 
the first 3 h, 100 pet of the mercury was 
eluted with 2N H2SO4. No gold and only 
about 10 pet of the silver was eluted 
with the mercury; however, about 20 pet 

■^Reference to specific products does 
not imply endorsement by the Bureau of 
Mines. 



of the silver was eluted with the mercury 
using the SBR resin. A subsequent elu- 
tion with 200 g/L NaCl in IN HCl removed 
all of the remaining silver from the five 
resins in 6 h with less than 10 pet of 
the gold being eluted. A final elution 
with 0.75 pet NaClO in 150 g/L NaCl plus 
5 g/L NaOH removed 93 to 98 pet of the 
gold in 9 h. 



Pregnant 

leoch 

solution 

(Au, Ag, Hg) 



Loaded SBR 



Resin 

adsorption 

circuit 



Barren 

leach 

solution 



H2SO4- 



NaCI plus HCl 



NoCIO in 
NoCI plus NoOh' 



SBR 
(Au,Ag,Hg) 



SBR 
{Au, Ag) 



SBR 
(Au) 



■Hg 



-Ag 



-Au 



Barren SBR 

FIGURE 1. - Flowsheet for recovery of mercury, 
silver, and gold from a loaded strong-base exchange 
resin (SBR) by sequential elution. 



TABLE 1. - Sequential elution of mercury, silver, and gold from strong-base 
ion-exchange resins 





Cumulative 
elution 


Metal eluted (cumulative) , pet 


Eluant solution 


Amberlite 


Amberlite 


Dowex 


Dowex 


Dowex 




time, h 


IRA-430 


IRA- 9 00 


21K 


SMA-1 


SBR 




EG 


Ag 


Au 


Hg 


Ag 


Au 


Hg 


Ag 


Au 


Hg 


Ag 


Au 


Hg 


Ag 


Au 


2N H2SO4 


1 
2 


82 
98 










93 
92 










78 
99 


4 
6 






56 
87 


3 

5 




1 


62 
95 


18 
18 












3 


100 








100 


5 





100 


7 





100 


9 





100 


20 





200 g/L NaCl in 


































IN HCl 


4 
5 


100 
100 


62 
85 



2 


100 
100 


65 
89 






100 
100 


69 
88 






100 
100 


57 
84 


2 
2 


100 
100 


62 
83 


n 









6 


100 


95 


5 


100 


97 





100 


96 


1 


100 


93 


2 


100 


93 







7 


100 


98 


7 


100 


99 





100 


98 


2 


100 


96 


5 


100 


97 


3 




8 


100 


100 


8 


100 


100 





100 


100 


4 


100 


98 


5 


100 


99 


3 




9 


100 


100 


8 


100 


100 





100 


100 


4 


100 


99 


5 


100 


100 


3 


0.75 pet NaClO in 


































150 g/L NaCl plus 


































5 g/L NaOH 


10 


100 


100 


75 


100 


100 


83 


100 


100 


31 


100 


100 


28 


100 


100 


68 




11 


100 


100 


83 


100 


100 


91 


100 


100 


58 


100 


100 


56 


100 


100 


79 




12 


100 


100 


87 


100 


100 


95 


100 


100 


79 


100 


100 


67 


100 


100 


87 




13 


100 


100 


87 


100 


100 


95 


100 


100 


91 


100 


100 


76 


100 


100 


90 




14 


100 


100 


88 


100 


100 


97 


100 


100 


93 


100 


100 


85 


100 


100 


91 




15 


100 


100 


89 


100 


100 


97 


100 


100 


95 


100 


100 


92 


100 


100 


92 




16 


100 


100 


91 


100 


100 


97 


100 


100 


96 


100 


100 


98 


100 


100 


94 




17 


100 


100 


92 


100 


100 


97 


100 


100 


96 


100 


100 


98 


100 


100 


94 




18 


100 


100 


93 


100 


100 


98 


100 


100 


97 


100 


100 


98 


100 


100 


95 



WEAK-BASE RESINS 



In addition to the work on sequential 
elution of strong-base resins, research 
was conducted with weak-base resins to 
determine their effectiveness in adsorb- 
ing gold and/or mercury from a caustic- 
cyanide solution. Both commercially 
available and experimental anion-exchange 
resins were examined. Experimental 



resins were included in the investigation 
because they effectively adsorb metal- 
cyanide complexes in the pH range in 
which most gold mills operate, pH 10 to 
11. Table 2 lists the weak-base resins 
tested in this investigation, along with 
available information concerning the res- 
in matrix material and functional group. 



TABLE 2. - Weak-base ion-exchange resins used in adsorption-elution 
experiments, by manufacturer 



Resin 



Sybron Corp. : A-305 

Diamond Shamrock Chemical Co: 

A-7 , 

A-340 , 

A-561 , 

Dow Chemical Co, : 

MWA-1 , 

WGR-2 , 

XFS-40114 , 

XFS-43309 

XFS-43356 , 

XU-40138 

XU-40139 



Functional group 



Poly amine, 



Secondary amine, 

Polyamine , 

...do , 



Dimethyl tertiary amine. 

Polyamine 

...do , 



NA. 
NA. 
NA. 
NA. 



Matrix 



Epoxy amine. 

Phenof ormaldehyde. 
Epoxy amine, 
Phenof ormaldehyde. 

Styrene-DVB. 
Epoxy amine. 
Do. 

NA. 
NA. 
NA. 
NA. 



NA Proprietary data not available from manufacturer. 



Batch-contact equilibrivim experiments 
were performed with each Ion-exchange 
resin to determine the effect of pH on 
the resin adsorption of gold and/or mer- 
cury. Six 1-g samples of each resin were 
equilibrated at a specific pH value be- 
tween 5 and 13 with NaOH and HCl. A syn- 
thetic feed solution (200 mL) contain- 
ing 500 ppm Au, 500 ppm Hg, and 0.5 g/L 
NaCN was placed in a 250-mL plastic di- 
gestion vessel, and the solution pH was 
adjusted to the respective pH of each 
resin. Resin was added and the mixture 
was rolled for about 48 h. Following the 
equilibrium contact, a sample of solution 
was removed for gold and mercury analy- 
sis, and the final equilibrium pH was 
measured. 

Figures 2 and 3 show the adsorption 
results for gold and mercury, respective- 
ly, for the commercially available res- 
ins A-305, A-7, A-340, A-561, MWA-1, and 
WGR-2 (as listed in table 2). The slopes 
of these curves are very typical for 
weak-base resins and show clearly the 
influence of pH on a resin's ability to 
adsorb metal-cyanide complexes. Better 
adsorption occurred in the pH range below 
9 for both gold and mercury, with none 
of the resins selective for gold over 
mercury. However, the A-305 resin indi- 
cated a preference for loading mercury 
over gold. 

Additional research performed on the 
A-305 resin indicated that this resin may 
be suitable for application in adsorbing 
mercury from existing gold operations. 
Testing was expanded with a mill leach 
solution to determine the effect of 
adsorption-elution recycling using a 
flow-through column and an experimental 
procedure. Figure 4 presents typical ad- 
sorption curves for both gold and mercury 
using this resin. Initially, gold ad- 
sorbed onto the resin, then quickly de- 
sorbed into the solution. Mercury, how- 
ever, steadily adsorbed. During the 
recycling experiments, the data showed 
that the resin's effectiveness to adsorb 
mercury decreased with each successive 
loading cycle. Attempts to regenerate 
the resin were unsuccessful. 



Subsequently, several experimental res- 
ins were tested using the batch-contact 
equilibrium technique previously de- 
scribed. Figures 5 and 6 show the ad- 
sorption results for gold and mercury, 
respectively, for a group of experimental 
resins developed by Dow Chemical Co. 
(XFS-40114, XFS-43309, XFS-43356, XU- 
40138, and XU-40139). Of particular in- 
terest are the three resins XFS-40114, 
XU-40138, and XU-40139. Gold and mercury 
were more efficiently adsorbed by these 
three resins, at a pH between 9 and 11, 
than they were by any of the other previ- 
ously examined resins. Also, the slopes 
of the curves for these resins were much 
steeper between pH 11 and 12, indicating 
that the resins should be easily eluted 
by a solution with a high pH value. Fol- 
lowing the adsorption tests, a sample of 
XFS-40114 was eluted using IM NaOH at am- 
bient temperature. Nearly 100 pet of the 
gold and 50 pet of the adsorbed mercury 
were recovered with 100 mL of eluant. 

Testing was expanded to determine the 
effectiveness of the three promising res- 
ins with actual leach solutions in a 
flow-through contact column. The leach 
solution contained 1.7 ppm Au, 3.0 ppm 
Ag, and 0.3 ppm Hg, with a pH of approxi- 
mately 10.5 and a free cyanide concentra- 
tion of 0.5 g/L. Approximately 5 L of 
solution was pumped through 1 g of resin 
in a 1-cm-ID glass column at a flow rate 
of 0.5 bed volume per minute. Following 
the adsorption phase, the loaded resin 
was eluted with 200 mL of IM NaOH at a 
flow rate of 0.17 bed volume per minute. 
Adsorption curves for gold, silver, and 
mercury are shown in figures 7, 8, and 9, 
respectively. The total adsorption and 
elution recoveries for these experiments 
are given in table 3. 

TABLE 3. - Summary of adsorption-elution 
results for selected resins, pet 



Resin 


Metal adsorption 


Metal eluted 




Au 


Hg 


Ag 


Au 


Hg 


Ag 


XFS-40114 

XU-40138 

XU-40139 


54.2 
57.7 
81.1 


51.8 
57.4 
66.6 


4.8 
10.8 
13.2 


13.7 
53.1 
13.1 




8.7 

6.0 


76.6 
85.1 
51.9 




8 9 10 

EQUILIBRIUM pH 

FIGURE 2. - Equilibrium adsorption curves for gold with commercially available exchange resins. 
(Broken curve patterns used for visual distinction only.) 




8 9 10 

EQUILIBRIUM pH 

FIGURE 3..- Equilibrium adsorption curves for mercury with commercially available exchange resins. 
(Broken curve patterns used for visual distinction only.) 



a. 

CL 



a. 

Q- 

<] 



Q. 
CC 
O 
(f) 
Q 
< 

_l 
UJ 




0.4 0.8 1.2 

SOLUTION VOLUME, L 

FIGURE 4. - Gold and mercury adsorption curves 
using resin A-305 with actual leach solution. 



100 



80 



2 60 

I- 

Q. 
CC 
O 
if) 

§ 40 



o 



20 - 







1 'C::::;^-^..^^ ' 


1 1 

KEY 


°\ ^^^^Sv\ 


XFS-43356 


\ ^nN/\ 


□ XFS-43309 


- h^ n\/\ 


A XFS-401 14 - 


\ ^vv ^ 


. ■ XU-40138 


\ ^k 


\« XU-40 139 


\ ^\ 




\ 


Ik^N. 


\ 


\\\ 


\ 


\ ^v\ 


«v 


V "^ 


\ 


\ ^\v 


\ \ 


\ IS. 


s ^ 


\ \\. 




-a. \_AcN° 


"^^ 


\ \^^ 


~-~-o 


^-— dv ^ 


1 1 1 


1 1 



7 



12 



8 9 10 II 

EQUILIBRIUM pH 

FIGURE 5. - Equilibrium adsorption curves for 
gold with experimental exchange resins. 



13 



100 



80 



60 - 



40 - 



20 







. ' "^ 


1 


1 




\ 


t^ 


\\ 








\ 


\X 








- ^\ 


\?v 






— 


\\ 




\ 






\\ 


V 


\\ 






- \ 


1 


\\\ 








w ^ 


\ 




\\ 




\\ 


\ 




\\ 




v\ 




\. 


KEY \ 
o XFS-43356 \> 




\ 




P 


■i 


N^ 


^ 


y 


Q XFS-43309 


^^. 


oS 


^ 




-A XFS-401 1 4 
■ XU-40138 


D 


-=rtr- 


\ 


^^^ 


• XU-40139 

1 1 


1 


1 




1 



8 



9 10 II 12 

EQUILIBRIUM pH 

FIGURE 6. - Equilibrium adsorption curves for 
mercury with experimental exchange resins. 



13 



CL 
CL 



CL 
CL 
<] 

o 

I- 

Q. 

q: 
o 
to 

Q 
< 



O 



uu 


^ 

f 


■ "h— ^^-Ili* 1 


' 




1 


80 


\ 




• 


\ 


— 


60 


- 


^^ 


=^^- 




^v • 






KEY 


^^^ 


^^ 


• N. 






A XFS-401 14 




"^ 


IT"^^ ■ 1 


40 
on 




■ XU-40138 
• XU-40139 

1 1 


1 




1 



I 2 3 

SOLUTION VOLUME, 



L 



FIGURE 7. - Gold adsorption curves for resins 
XFS-40n4, XU-40138, and XU-40139 with actual 
leach solution. 



o 

Q. 



CL 
Q. 



80 



60 



Q. 
CL 

< 40 



a. 
cr 
o 

CO 
Q 

< 
q: 

LU 
CO 



20 








KEY 
A XFS-40 1 14 
■ XU-40138 
• XU-40139 



00 



12 3 4 

SOLUTION VOLUME, L 

FIGURE 8. - Silver adsorption curves for resins 
XFS-401 14, XU-40138, and XU-40139 with actual 
leach solution. 



Q- 




Ci. 


80 


^ 




a 




Q. 




< 




o 


60 


1- 




LL 




cr 




o 




CO 




< 


40 


>- 




en 




3 




o 




cr 

LU 


20 



Vn. 


1 1 


1 1 

KEY 


^v^ 




A XFS-401 14 


-> 


^^ 


■ XU-40138 
• XU-40139 


— 




K^ •\« • 


^^^^ 




1 1 


■ 
1 1 



12 3 4 

SOLUTION VOLUME, L 

FIGURE 9. - Mercury adsorption curves for resins 
XFS-40114, XU-40138, and XU-40139 with actual 
leach solution. 



These three resins (XFS-40114, XU- 
40138, and XU-40139) were generally se- 
lective for gold and mercury, with mini- 
mal silver and other base metals being 
adsorbed. The stronger the attraction a 
resin had for gold and mercury, the less 
selective it was and the poorer the elu- 
tion with NaOH. The exchange resin XU- 
40138 appeared to show the best loading 
and eluting characteristics. During this 



experiment, XU-40138 loaded to about 140 
tr oz/st Au and produced an eluant con- 
taining an average of 12.4 ppm Au. By 
comparison, XU-40139 and XFS-40114 loaded 
134 and 97 tr oz/st Au, respectively, and 
eluted average Au concentrations of 1.7 
and 3.9 ppm, respectively. The weak-base 
resin XU-40138 may have a commercial ap- 
plication if cyclic loading-elution tests 
are successful. 



SUMMARY AND CONCLUSIONS 



Strong-base ani on-exchange resins were 
loaded with gold, silver, and mercury, 
then sequentially eluted with various 
eluants. Sequential elution of Amberlite 
IRA-430, for example, eluted 100 pet of 
the mercury with 2N H2SO4, 100 pet of the 
silver with 200 g/L NaCl in IN HCl, and 
all of the remaining gold with 0.75 pet 
NaClO in 150 g/L NaCl plus 5 g/L NaOH. 
Depending on the strong-base resin exam- 
ined, this elution scheme allowed little 
contamination between the mercury and the 
precious metals. 



In addition to strong-base resins, 
weak-base resins were investigated for 
possible applications in the caustic- 
cyanide systems of the gold industry. 
None of the commercial weak-base resins 
tested was suitable for a caustic-cyanide 
system; however, three experimental res- 
ins have possible application because of 
their effectiveness in the pH range of 
10 to 12. When using a mill leach solu- 
tion, these resins adsorbed the gold and 
mercury with only minimal silver being 
adsorbed. 



REFEEIENCES 



1. Arizona Pay Dirt. Gold Fields De- 
velops State of the Art Leaching System 
at Ortiz. No. 516, 1982, pp. 41-42, 44, 
46, 48. 

2. Engineering and Mining Journal. 
Alligator Ridge Uses Heap Leaching To 
Produce Gold Bullion Bars. V. 182, No. 
8, 1981, pp. 35, 37. 

3. Grace, K. A. Exploration and De- 
velopment in 1981. World Min. , v. 183, 
No. 7, 1981, pp. 58-62. 

4. Jackson, A. Jerrit Canyon Project. 
Eng. and Min. J., v. 183, No. 7, 1982, 
pp. 54-58. 

5. Skillings, D. N. , Jr. Getty Mining 
Co. Starting Up Its Mercury Gold Opera- 
tion in Utah. Skillings' Min. Rev., v. 
72, No. 17, 1983, pp. 4-9. 

6. . Homes take Proceeding With 

Its McLaughlin Gold Project. Skillings' 
Min. Rev., v. 72, No. 4, 1983, pp. 3-6. 

7. . Pinson Mining Co. Mark- 
ing First Full Year of Gold Production. 
Skillings' Min. Rev., v. 71, No. 28, 
1982, pp. 8-12. 

8. . Smokey Valley Operations at 

Its Round Mountain Mine in Nevada. 
Skillings' Min. Rev., v. 68, No. 9, 1979, 
p. 8. 

9. Steel, G. L. Candelaria: Famous 
Silver Producer. Min. Eng. (Littleton, 
CO), V. 33, No. 6, 1981, pp. 659-660. 



10. Laxen, P. A., G. S. M. Becker, and 
R. Rubin. Developments in the Applica- 
tion of Carbon-in-Pulp to Recovery of 
Gold From South African Ores. J. S. Afr. 
Inst. Min. & Metall. , v. 79, No. 11, 
1979, pp. 315-326. 

11. Potter, G. M., and H. B. Salis- 
bury. Innovations in Gold Metallurgy. 
(Pres. at Am. Min. Congr. Min. Conv. and 
Environ. Show, Denver, CO, Sept. 9-12, 
1973.) BuMines preprint (Salt Lake City, 
UT), 1973, 12 pp. 

12. Hussey, S. J. Application of Ion 
Exchange Resins in the Cyanidation of a 
Gold and Silver Ore. BuMines RI 4374, 
1949, 34 pp. 

13. Burstall, F. H. , P. J. Forrest, 
N. F. Kember, and R. A, Wells. Ion Ex- 
change Process for Recovery of Gold From 
Cyanide Solution. Ind. and Eng. Chem. , 
V. 45, No. 8, 1953, pp. 1648-1658. 

14. Mooiman, M. D. , J. 0. Miller, 
J. B. Hiskey, and A. R. Hendriksz. Com- 
parison of Process Alternatives for Gold 
Recovery From Cyanide Leach Solutions. 
Paper in Proc. Soc. Min. Eng. AIME Fall 
Meeting (Salt Lake City, UT, Oct. 19-21, 
1983). AIME, 1984, pp. 93-107. 

15. Von Michaelis, H. Innovation in 
Gold and Silver Recovery. Soc. Min. Eng. 
AIME, preprint 83-119, 1983, 9 pp. 



10 



STAGED HEAP LEACHING-DIRECT ELECTROWINNING 
By C. H. Elgesi and M. D. Wroblewski2 



ABSTRACT 



The Bureau of Mines is conducting re- 
search to develop a staged, agglomeration 
heap leaching system employing direct 
electrowinning of dissolved precious met- 
als from the pregnant leaching solution. 
This paper describes bench-scale research 
the Bureau conducted to develop such a 
system and an efficient electrowinning 
cell. 

The Bureau demonstrated through leach- 
ing simulations that pregnant solutions 



suitable for direct electrowinning can be 
generated by staged heap leaching. The 
Bureau also demonstrated that improved 
mass transfer (IMT) cells can recover 
precious metals from low-grade pregnant 
solutions generated by cyanide heap 
leaching. This research indicates that 
staged heap leaching-direct electrowin- 
ning is a promising technique for recov- 
ering precious metals from low-grade 
resources. 



INTRODUCTION 



Heap leaching has been used to recover 
gold and silver from low-grade ores since 
the early 1970' s, and numerous operations 
in the Western United States are present- 
ly using heap leaching to produce pre- 
cious metals (J:.~3.)»'^ One stimulus for 
the rapid increase in the number of com- 
mercial operations has been the dramatic 
and sustained increase in the price of 
gold. Another stimulus has been a suc- 
cession of improvements in the practice 
of heap leaching that has made possible 
the treatment of increasingly difficult 
ores. 

Current heap leaching practice for pre- 
cious metals includes sprinkling di- 
lute alkaline NaCN solution on heaps of 
crushed material, which may have been 
agglomerated with cement or lime (4-^) , 
and allowing the solution to percolate 
through the heap. The effluent is col- 
lected and passed through activated car- 
bon beds that adsorb the precious metals. 
For ores in which the silver content 
is high relative to the gold content, 
Merrill-Crowe zinc precipitation may be 

^Chemical engineer. 

^Physical science technician. 
Reno Research Center, Bureau of Mines, 
Reno, NV. 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



favorable ; this technique avoids the 
large carbon inventories required to ad- 
sorb the larger quantity of silver. When 
carbon is used, the precious metals are 
removed by stripping at elevated tempera- 
ture with a strong caustic-cyanide solu- 
tion, often with the addition of methyl 
or ethyl alcohol (7). Gold and silver 
are recovered from the stripping solution 
by electrowinning onto a steel wool cath- 
ode and are then fire-refined to produce 
dore metal. 

The carbon adsorption-desorption step 
is an efficient way to concentrate the 
typically dilute gold solutions (about 
1 mg/L) from heap leaching. However, be- 
cause the carbon is usually acid-washed 
and must be thermally regenerated before 
reuse, the carbon adsorption-desorption 
step can add significantly to the operat- 
ing and capital costs of a heap leaching 
operation. In marginal cases, this step 
may make an operation uneconomical. In 
addition, there are precious metals 
losses associated with loss of carbon 
from the system. For these reasons, it 
is desirable to eliminate the carbon 
adsorption-desorption step and directly 
electrowin the gold from the heap 
effluent. 

For staged heap leaching-direct elec- 
trowinning to be efficient, the precious 
metals concentration of the pregnant so- 
lution generated during leaching must be 



11 



higher than that obtained by conventional 
heap leaching, and the solutions may have 
to be fortified with electrolyte. The 
precious metals concentration of the 
pregnant solution can be increased by 
(1) leaching with smaller volumes of so- 
lution, (2) agglomeration pretreatment 
with cyanide solution (which starts the 
leaching process during the curing peri- 
od) , and (3) cycling the leaching solu- 
tion through more than one heap in a 
staged manner. The resultant pregnant 
solution is passed directly through an 
electrowinning circuit. The barren solu- 
tion from electrowinning is recycled to 
the leaching circuit. 

For direct electrowinning of precious 
metals to be commercially feasible, it 
is necessary to obtain suitable recov- 
eries in electrolytic cells of reason- 
able size and at ambient temperature, be- 
cause heating large volumes of solution 
is costly. This objective is made more 



difficult because the concentration of 
precious metals in heap leaching solu- 
tions is very low, A typical heap leach- 
ing pregnant solution contains 0,5 to 2.0 
mg/L Au, compared to 50 to 1,000 mg/L Au 
in carbon stripping solutions from which 
gold is conventionally electrowon. The 
requirements for direct electrowinning 
are (1) precious metals electrowinning 
cells that operate more efficiently at 
ambient temperature than those currently 
in use and (2) a method for conducting 
heap leaching operations that maximizes 
the precious metals content. The Bureau 
is currently investigating direct elec- 
trowinning of gold and silver from heap 
leaching solutions and has developed 
electrolytic cells with IMT characteris- 
tics. Countercurrent staged heap leach- 
ing techniques are being improved in 
order to increase the concentration of 
precious metals in the heap effluent. 



DIRECT ELECTROWINNING 



CELL DESIGN 

The Bureau's original IMT cell design 
is similar to that of the Zadra cell, 
which has been the industry standard for 
gold electrowinning for many years (8^) . 
The Zadra cell (fig. 1) features three 
concentric circular containers. The in- 
ner container, which serves as the cath- 
ode compartment, is a perforated insula- 
tor that contains a central feed tube, 
current distributor, and steel wool, onto 
which precious metals are deposited. The 



Pregnant solution • 




CROSS SECTION 

FIGURE 1. - Zadra cell design 



Barren solution 
SIDE VIEW 



anode, a circular stainless steel screen, 
is outside the cathode in the second con- 
tainer. Pregnant solution enters through 
the central feed tube and flows upward 
and outward through the steel wool, 
Zadra-type cells have been used to elec- 
trowin precious metals from stripping 
solutions at temperatures of 70° to 
85° C. However, several disadvantages of 
the Zadra design are that solution flow 
is unevenly distributed, "effective" 
electrode spacing is excessive,^ and cell 
volume is utilized inefficiently. 

The circular IMT cell (fig. 2) was 
designed to overcome the problems of the 
Zadra cell by providing uniform solution 
flow across and through the cathodic 
steel wool. The solution distribution 
tube is stainless steel and serves as a 
second anode, which decreases the effec- 
tive electrode spacing. To increase the 
cell efficiency, provision was made for 
rapid internal solution circulation. Al- 
though the mean residence time is unaf- 
fected by this internal circulation, the 

^The physical properties of the pervi- 
ous steel wool cathode preclude exact 
measurement of electrode spacing. 



12 



Pregnant solution 



Recirculation pump 



>> 



Anode 



n> 



® 



ffi] 



f3S;i 









© 



.Cathode 



h 



L 



■Barren solution 



Cathode compartment 
(steel wool) 

Anode compartment 



FIGURE 2. - Design of large IMT cell. 

highly turbulent flow results in a thin- 
ner electrode boundary layer, decreases 
concentration polarization, and increases 
metal deposition rates up to 200 pet. 
Internal circulation flow rates were 10 
to 25 times the pregnant solution feed 
rate. The circular cell was built in 
several sizes, ranging from 55 to 888 
cm^ (cathode volume). 

The data presented in this paper were 
generated using circular IMT cells, al- 
though current research efforts are con- 
centrated on a rectangular IMT cell. The 
rectangular cell is similar in appearance 
to "breadbasket" types currently in use. 
It should scale up easily while retaining 
the favorable operating characteristics 
of the circular cell, 

ELECTROWINNING TESTS 

Direct electrowinning tests were made 
using small (55 cm^ cathode volume) and 
large (888 cm^) circular IMT cells. The 
cells were operated at ambient tem- 
perature (20* to 25° C) with voltage, 
amperage, feed rate, internal circula- 
tion rate, NaOH concentration, and pre- 
cious metals concentration as variables. 
Small-cell tests were performed by pump- 
ing solution from a reservoir through the 
cell numerous times , using various flow 
rates and operating times. Large-cell 
tests were made by passing feed solution 
through the cell once, but with internal 
circulation of the solution. 

For comparison, direct electrowinning 
tests were made using small (67 cm^ cath- 
ode volume) and large (888 cm^) Zadra 




0.05 0.10 0.15 

NaOH CONG, M 



0.20 



FIGURE 3. - 
concentration. 
72 mL ^min.) 



Recovery of gold with increasing NaOH 
(Cell operated at 3 V and feed rate of 



cells. The cells were operated at 3 V 
and ambient temperature (20° to 25° C) , 
Retention time in the cells was varied, 

RESULTS AND DISCUSSION 

Operating data and results are given 
for the small circular IMT cell in table 
1 and for the large circular IMT cell in 
tables 2 and 3, The results show that 
flow rate and NaOH concentration are im- 
portant parameters. Pregnant solutions 
from heap leaching are unsuitable as cell 
electrolytes unless they are fortified 
with electrolyte. Figure 3 shows the in- 
crease in gold recovery with the addition 
of NaOH to 0,05M (4 Ib/st of solution) in 
the large IMT cell. 

Although NaOH was the most effective 
additive for fortifying the electrolyte, 
other salts were investigated, Na2C03 
functioned well, but was not as effective 
on a molar basis. Two salts of strong 
acids, Na2S0i+ and NaN03 , were also inves- 
tigated but caused severe corrosion of 
the stainless steel anodes. 

The effects of the circulation flow 
rate on gold and silver recovery in the 
small IMT cell are shown in figure 4, 
Recovery of both metals increased mark- 
edly with increased flow rates up 
to approximately 250 mL/min and then 
decreased. 

The effects of the internal circulation 
flow and feed rates on gold and silver 



13 



TABLE 1. - Performance of small IMT celll 



Current , 


Operating 
time, h 


Flow rate, 
mL/min 


NaOH cone, 
M 


Recovery, pet 


Deposition i 


rate, mg/min 


A 


Au 


Ag 


Au 


Ag 


0.07 


1.0 


30 


0.10 


54 


73 


0.3 


0.6 


.08 


1.0 


75 


.10 


74 


81 


.5 


.7 


.08 


1.0 


150 


.10 


90 


96 


.6 


.8 


.09 


1.0 


240 


.10 


92 


100 


.8 


1.0 


.09 


1.0 


420 


.10 


84 


99 


.6 


.9 


.09 


1.5 


250 


.10 


97 


100 


.5 


.7 


.03 


2.0 


250 


.01 


43 


26 


.2 


.1 


.03 


2.0 


250 


.02 


46 


31 


.2 


.1 


.06 


2.0 


250 


.05 


61 


78 


.2 


.4 


.08 


2.0 


250 


.10 


84 


91 


.3 


.4 


.09 


2.0 


250 


.15 


92 


92 


.3 


.4 


.12 


2.0 


250 


.20 


96 


96 


.3 


.5 



^Cell operated at 3 V. Pregnant synthetic solutions contained approximately 40 
mg/L Au and 50 mg/L Ag. Cathode packing density, 0.018 g/cm^ (1 g steel wool). 

TABLE 2. - Performance of large IMT cell showing effects of internal 
circulation flow rate and feed rate^ 



Current, 


Retention 


Internal circulation 


Feed rate, 


Recovery, 


Deposition rate. 


A 


time, min 


flow rate, L/min 


mL/min 


pet 


mg/min 




Au 


Ag 


Au 


Ag 


0.6 


8.6 





250 


43 


56 


4.5 


7.0 


.7 


8.6 


.5 


250 


60 


70 


6.3 


8.8 


.7 


8.6 


1.0 


250 


67 


78 


6.5 


10.3 


.8 


8.6 


1.5 


250 


73 


85 


6.8 


10.2 


.8 


8.6 


2.0 


250 


75 


81 


7.4 


9.1 


.8 


21.5 


2.0 


100 


86 


91 


3.7 


6.4 


.8 


4.3 


2.0 


500 


51 


63 


8.1 


16.0 



^Cell operated at 3 V with O.IM NaOH. Pregnant synthetic solutions contained 
approximately 40 mg/L Au and 50 mg/L Ag. Cathode packing density, 0.018 g/cm^ (16 
g steel wool) . 

TABLE 3. - Performance of large IMT cell showing effects of cell voltage 
and NaOH concentration^ 



Potential, 


Current, 


NaOH cone. 


PH 


Au recov- 


Deposition 


Current efficiency. 


V 


A ■ 


M 




ery, pet 


rate, mg/min 


pet 


1.5 


~o 


0.05 


12.75 











2.0 


.1 


.05 


12.75 


39 


1.8 


23.4 


2.5 


.3 


.05 


12.75 


64 


2.0 


12.9 


3.0 


.4 


.05 


12.75 


90 


2.8 


14.8 


3.5 


.6 


.05 


12.75 


95 


2.9 


9.8 


4.0 


.8 


.05 


12.75 


97 


3.0 


7.6 


3.0 


.1 


~0 


9.35 


12 


.4 


7.3 


3.0 


.2 


.01 


10.45 


68 


2.1 


21.0 


3.0 


.4 


.05 


12.75 


90 


2.8 


14.8 


3.0 


.5 


.10 


13.05 


94 


2.9 


11.5 


3.0 


.7 


.15 


13.13 


96 


3.0 


8.5 


3.0 


1.0 


.20 


13.40 


97 


3.0 


6.3 



^All tests used a retention time of 12.3 min, a feed rate of 72 mL/min, and an in- 
ternal circulation flow rate of 2 L/min. Pregnant solutions contained 43 mg/L Au and 
1 mg/L Ag. Cathode packing density, 0.018 g/cm^. 



14 



100 
^ 90 


/ 


/C^-^ 


^Silver 


o 


/ 


X 


^^^^^ _ 


Q. 


/ / 


^r 


^© 


.80 

>- 


'// 


^^Gold 




uj 70 


>/ 






> 


/ 






8 60 


-/ 






UJ 


1 






^50 


w 






40 


1 


1 1 


1 1 



100 200 300 400 500 

FLOW RATE, mL/min 

FIGURE 4. - Recovery of gold and silver with in- 
creasing circulation flow rate after 1 h of operating 
time. (O.OlMNaOH.) 




2 3 4 5 

CELL POTENTIAL, V 

FIGURE 5. - Recovery of gold with increasing cell 
potential. (0.05M NaOH; feed rote, 72 mL/min; in- 
ternal circulation flow rate, 2 L/min; retention time, 
30 min.) 



recovery in the large IMT cell are shovm 
in table 2. Gold and silver recovery and 
deposition rates increased as the inter- 
nal circulation flow rate increased. 
When the feed rate was increased, gold 
and silver recovery decreased, but the 
deposition rates increased. The data 
show that silver is more easily electro- 
won than gold; consequently, adjustment 
of electrowinning parameters to achieve 
the best gold recovery will ensure good 
silver recovery. 

The effects of decreasing gold con- 
centration in the feed to the small 
IMT cell are shown in table 4, Deposi- 
tion rates decreased with decreasing 
gold concentration, and the current effi- 
ciency decreased to <1 pet when the gold 



concentration in the pregnant solution 
was less than 7 mg/L, 

The effects of cell voltage on gold re- 
covery and cell current in the large IMT 
cell are shown in figure 5, Current var- 
ied directly with applied voltage, and 
gold recovery increased at greater cell 
currents. Figure 6 shows that while gold 
deposition rates increased with increased 
cell potential, current efficiency de- 
creased sharply. When the cell potential 
was increased above 2 V, a greater pro- 
portion of the current was used in the 
decomposition of water as opposed to the 
deposition of gold. 

Table 5, which compares the opera- 
tion of large Zadra and IMT cells, shows 
that gold and silver recovery, current 



TABLE 4. - Performance of small IMT cell with decreasing gold concentration 
in feed solution^ 



Au in 


pregnant sol, 
mg/L 


Au recovered in 1 h of 
operating time, pet 


Deposition rate, 
mg/min 


Current efficiency, 
pet 


48 


92 
93 
91 
84 
82 
85 
81 


0.72 
.60 
.28 
.11 
.07 
.03 
.015 


9.6 


37 


8.6 


10 


2.3 


7.4 


1.5 


3.8 


.71 


2.1 


.50 


1.1 


.20 



^Cell operated at 3 V, with a flow rate of 240 mL/min, a retention time of 3.3 min, 
an NaOH concentration of O.IOM, and a cathode packing density of 0.018 g/cm^. 



TABLE 5. - Recovery of gold and silver in large electrolytic cells ^ 



Feed rate, 
mL/min 



Current , 

A 



Synthetic 
pregnant 
solu t ion , mg 



Au 



Ag 



Recovery, mg 



Au 



Ag 



Current 

efficiency, 

pet 



Au 



Ag 



Deposition 
rate, 
mg/min 



Au 



Ag 



ZADRA CELL 



100 


0.60 


592 


720 


290 


403 


2.5 


6.3 


1.8 


2.5 


250 


.75 


656 


880 


112 


336 


1.9 


10.5 


1.8 


5.3 


500 


.75 


640 


896 


64 


208 


2.2 


12.9 


2.0 


6.5 



IMT CELL 



100 


0.95 


688 


1,136 


594 


1,030 


3.2 


10.1 


3.7 


6.4 


250 


.95 


608 


816 


440 


629 


5.9 


15.4 


6.9 


9.8 


500 


.97 


512 


816 


261 


512 


6.9 


24.6 


8.2 


16.0 



•'^Both cells were operated at ambient temperature, at 3 V, with a cath- 
ode packing density of 0.018 g/cm^ , a feed volume of 16 L, and O.IM NaOH. 
The IMT cell had an internal circulation flow rate of 2 L/min. Recovery 
based on 1 pass through cell. 



15 



efficiency, and deposition rate, were all 
substantially higher for the IMT cell 



than for the Zadra cell , 
higher feed rates. 



especially at 



STAGED HEAP LEACHING 



HEAP PREPARATION AND LEACHING CYCLE 

Staged heap leaching was developed to 
produce pregnant solutions with the high- 
est possible gold concentration. Figure 
7 is a schematic of staged heap leaching 
operated in conjunction with direct elec- 
trowinning. Using this scheme, several 
heaps are leached countercurrently be- 
fore the pregnant solution is routed to 



0.20 



llJ 


$ 


1- 


Q> 


< 


0) 


a: 


(0 




».— 


z 


o 


O E 




m 


H 


C3> 


(D 


h- 


O 


K 


CL 


r 


Q 






O) 




E 




.05- 



2 3 4 
CELL POTENTIAL, V 

FIGURE 6. - Effect of cell potential on deposition 
rate and current efficiency. (O.OSM^ NaOH; feed rate, 
72 mL/min; internal circulation flow rate, 2 L/min, 
retention time 30 min.) 



electrowinning. The ore in the heaps is 
agglomerated with cement and a strong 
NaCN solution. The leaching action of 
NaCN starts during the curing period and 
may be almost finished by the time the 
heap is constructed. The dissolved Au 
and Ag can be removed from the heap by 
repeated washing with a small volume of 
water or dilute NaCN solution. Leaching 
solution is applied intermittently be- 
cause a "pulsed" flooding cycle resulted 
in higher precious metals extraction and 
used less leachant. 



Final wash 



NaCN ■ 
Cement • 




Agglomeration 



Next heap 
to be leached 

FIGURE 7. - Staged heap leaching-direct 

electrowinning. 



16 



After leaching of a heap is completed, 
a thorough washing cycle is conducted to 
recover additional values. The wash wa- 
ter is fortified with NaCN and used to 
agglomerate a new charge of ore for heap 
leaching. 

LABORATORY TESTS 

To conduct staged heap leaching-direct 
electrowinning experiments in the labora- 
tory, 22.7-kg (50-lb) charges of ore were 
agglomerated and percolation leached in 
acrylic columns 5-ft high by 5.5 in ID. 
Agglomeration was accomplished with a 
disk pelletizer in which the 22.7-kg ore 
charges were combined with 227 g portland 
cement (20 Ib/st ore) and 2.5 L of 0.1 OM 
NaCN-0.05M NaOH solution (9.8 and 4.0 lb/ 
st solution, respectively). The agglom- 
erated ore was placed in columns and aged 
for at least 24 h. A small IMT cell con- 
taining 1 g of steel wool and operated at 
3 V and a flow rate of 250 mL/min was 
used to recover gold and silver from 
solution. 

Each test series used five columns, 
each containing a 22.7-kg charge of ag- 
glomerated ore. As shown in figure 8, 
the laboratory procedure was to bring 
columns on-line in stages , such that 
steady-state operation could be approx- 
imated by the completion of the test 
series. In stage 1, 1 L of 0.05M NaOH 
(4 Ib/st solution) was used to start 
the leaching process; after progressing 
to stage 3, there were three 1-L batches 



of leachant in the system. Electrowin- 
ning was conducted after each cycle of 
leaching, and the barren cell electrolyte 
was recycled to the next stage of leach- 
ing. Additions of NaOH, when required, 
were made prior to electrowinning, while 
solution volumes were adjusted prior to 
each leaching step. Upon completion of 
leaching of columns 1 and 2, these col- 
umns were subjected to wash cycles using 
2.5-L quantities of water. The wash wa- 
ters were then fortified with NaCN and 
NaOH and used to agglomerate the ore 
charges to columns 4 and 5, respectively. 
Objectives throughout the leaching- 
electrowinning sequence were to obtain 
maximum precious metals recovery while 
maintaining the greatest possible pre- 
cious metals tenor in the pregnant leach 
solutions. Pregnant solution analyses 
are ideally based upon the average of the 
pregnant leaching solutions produced dur- 
ing stage 5. Precious metals recoveries 
are ideally based upon the tails analysis 
of column 3 after completion of stage 5. 
A total of 25 to 30 cycles of leaching- 
electrowinning were employed per test 
series. 

RESULTS AND DISCUSSION 

Results of staged heap leaching-direct 
electrowinning tests on four ores of dif- 
fering grade and mineralogy are given 
in table 6. The gold concentration of 
the pregnant solutions produced was a 
function of the ore grade. For the ores 



STAGE 1 



STAGE 2 



STAGE 3 



STAGE 4 



STAGE 5 







1 






L 


Electrowinning 





















1 






2 






L 


— 












Electrowinning 



Electrowinning 



ri 



Electrowinning 



n 



Electrowinning 



FIGURE 8. - Laboratory-scale staged leaching of agglomerated ores. (Numbers identify leaching 
columns.) 



17 



TABLE 6. - Recovery of gold by staged heap 
leaching-direct electrowinnlng 



Ore grade ppm Au.. 

Gold recovery from ore, pet: 

Staged heap leaching 

Conventional leaching 

Average content of pregnant 

solution mg/L. . 

NaOH used kg/mt ore. . 

Current efficiency pet . . 

Gold recovery, electrowinnlng, 
pet: 

Batch (average) 

Overall 



Test 



9.3 

87 

86 

32.5 

0.43 
7.3 



89 
98 



A. 8 

91 

99 

26.7 

0.49 
2.0 



90 
99 



2.1 

83 

83 

12.3 

0.51 
1.5 



91 
99 



1.1 

73 

74 
4.4 

0.41 
0.7 



90 
99 



tested, the pregnant solutions generated 
in each ease were at least an order of 
magnitude higher in precious metals val- 
ues than would be typical of solutions 
generated using conventional leaching 
practice. The increases in precious met- 
als tenor were achieved in three of the 
four tests with no measurable sacrifice 
in precious metals recovery. Overall re- 
covery of precious metals from solution 
by electrowinnlng was more than 98 pet in 
all eases. 



Impurity buildup was not a problem dur- 
ing the staged heap leaching-direct elec- 
trowinnlng tests. Impurities were close- 
ly monitored during test 4, which was 
conducted on gold ore containing 1.1 g/mt 
Au (0.035 tr oz/st Au) . Impurities that 
accumulated in the pregnant feed to the 
IMT cell were silicon (4 ppm) , mercury 
(10 ppm), calcium (20 ppm), copper (21 
ppm), and zinc (200 ppm). Only the mer- 
cury codeposited with the precious 
metals. 



CONCLUSIONS 



The research demonstrated that IMT 
cells can recover precious metals from 
low-grade pregnant solutions generated by 
cyanide heap leaching. Efficient opera- 
tion at ambient temperature makes the IMT 
cell suitable for direct electrowinnlng. 
Major differences in comparison with the 
standard Zadra cell are the rapid recir- 
culation of solution within the cell, 
better control of solution flow streams, 
and ambient temperature operation. 



Simulated staged heap leaching demon- 
strated that suitable pregnant solutions 
for direct electrowinnlng can be gener- 
ated. The buildup of impurities in the 
recycled leachant and/or electrolyte did 
not affect the electrowinnlng step in the 
number of cycles studied. Staged perco- 
lation leaching-direct electrowinnlng 
shows considerable promise as a technique 
for recovering precious metals from low- 
grade resources. 



18 



REFERENCES 



1. Chamberlain, P. C, and M. G. Po- 
jar. Gold and Silver Leaching Practices 
in the United States. BuMines IC 8969, 
1984, 47 pp. 

2. Eisele, J. A,, A. F. Colombo, and 
G. E, McClelland. Recovery of Gold and 
Silver From Ores by Hydrometallurgical 
Processing, Paper in Precious Metals: 
Mining, Extraction and Processing. 
(Proc. Int. Symp. held at AIME Annu. 
Meeting, Los Angeles, CA, Feb. 27-29, 
1984). AIME, 1984, pp. 387-395. 

3. McQuiston, F. W. , and R. S. Shoe- 
maker. Gold and Silver Cyanidation Plant 
Practice, v. 2. Metall. Soc. AIME, 1981, 
263 pp. 

4. Heinen, H. J., G. E. McClelland, 
and R. E, Lindstrom. Enhancing Perco- 
lation Rates in Heap Leaching of Gold- 
Silver Ores. BuMines RI 8388, 1979, 
20 pp. 



5. McClelland, G. E., and J. A. 
Eisele. Improvements in Heap Leaching 
To Recover Silver and Gold From Low- 
Grade Resources. BuMines RI 8612, 1982, 
26 pp. 

6. McClelland, G. E., D. L. Pool, 
and J. A. Eisele. Agglomeration-Heap 
Leaching Operations in the Precious Met- 
als Industry. BuMines IC 8945, 1983, 
16 pp. 

7. Heinen, H. J., D. G. Peterson, and 
R. E, Lindstrom. Processing Gold Ores 
Using Heap Leach-Carbon Adsorption Meth- 
ods. BuMines IC 8770, 1978, 21 pp. 

8. Zadra, J. B., A. L. Engel, and 
H. J. Heinen. Process for Recovering 
Gold and Silver From Activated Carbon 
by Leaching and Electrolysis. BuMines 
RI 4843, 1952, 32 pp. 



19 



MERCURY PRECIPITATION DURING CYANIDE LEACHING OF GOLD ORES 

By Richard G. Sandberg'' 

ABSTRACT 



Many gold-bearing ores throughout the 
Western United States contain small quan- 
tities of mercury. During cyanidation, 
10 to 40 pet of the mercury is extracted 
along with the precious metals. The 
presence of mercury decreases gold load- 
ing and increases stripping time on acti- 
vated carbon, complicates fire refining 
of the gold cathodes, and creates a pos- 
sible health hazard. In the investiga- 
tion described in this paper, the Bureau 
of Mines examined several methods for re- 
moving mercury from gold-silver cyanide 
leach slurries. 

CaS addition to cyanide leach slurries 
or to a laboratory ball mill containing 



NaCN and lime reduced mercury dissolution 
to < 0.5 pet. Mercury loading on acti- 
vated carbon was reduced to < 0.2 pet. 
Gold loading on activated carbon was af- 
fected very little by sulfide addition; 
however, silver loading was reduced to 
to 6 pet, as opposed to the typical val- 
ues of 90 to 100 pet of the silver and 
none of the mercury being adsorbed on the 
carbon. 

Preliminary testing in a mill operation 
using NaHS showed that mercury precipi- 
tation was nearly complete at the point 
of addition; however, as with Na2S, 30 
to 50 pet of the precipitated HgS redis- 
solved with time. 



INTRODUCTION 



Mercury contamination has been a prob- 
lem in the recovery of gold and silver 
from many western deposits, which may 
contain as much as 20 ppm Hg. During the 
cyanide leaching process, 10 to 40 pet of 
the mercury is normally solubilized along 
with gold and silver. The mercury must 
be recovered or precipitated so that it 
does not present a health hazard during 
electrolysis, smelting of the cathodes, 
and regeneration of activated carbon. 
Some gold mill operations recover mer- 
cury by retorting the cathodes prior to 
smelting (J^-2)^ or autoclaving the ore to 
extract minimal mercury (3). 

In an effort to reduce mercury solu- 
bilization, the Bureau of Mines conducted 
bench-scale leaching tests with an ore 
containing 0.08 tr oz/st Au, 0.06 tr oz/ 
St Ag, and 17 ppm Hg ( 4_) . Leaching was 
accomplished with cyanide in an air- 
agitated Pachuca-type vessel. Mercury 

^ Group supervisor, Salt Lake City Re- 
search Center, Bureau of Mines, Salt Lake 
City, UT. 

''Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



extraction was decreased from 40 to 10 
pet by (1) decreasing NaCN concentration 
from 20 to 0.34 lb per short ton of solu- 
tion, (2) decreasing the pH from 11,5 to 
11, and (3) increasing particle size from 
minus 270 mesh to minus 10 plus 48 mesh. 
Although mercury extraction was reduced, 
it was not eliminated. 

The remaining soluble mercury can be 
precipitated with sulfides as shown by 
the chemical reaction 

Hg(CN)^" + MS > HgS + m2+ + 4CN" , 

Calcium, sodium, silver, iron, and zinc 
sulfides have been used to precipitate 
mercury from Au(CN)3 solutions (^~5^) • 
Addition of Ag2S would tie up considera- 
ble silver, and excess Ag2S would have to 
be recovered. FeS is undesirable because 
of cyanide loss due to ferrocyanide for- 
mation. Calcium and sodium sulfides are 
effective in precipitating mercury and 
are not harmful to gold recovery. This 
paper reports on a number of tests in 
which calcium and sodium sulfides were 
used to remove mercury from gold and 
gold-silver cyanide leach slurries. 



20 



MERCURY PRECIPITATION FROM SLURRY 



Gold ore containing 17 ppm Hg was 
leached with solutions containing 0.34 or 

6 lb NaCN per short ton of solution and 
enough lime to give a pH of 11, tjien con- 
tacted with Na2S. The results of these 
tests are shown in figure 1. Increasing 
Na2S lowered mercury extraction when low 
concentrations of cyanide (0.34 Ib/st 
NaCN) were used. However, when high con- 
centrations of cyanide (6 Ib/st NaCN) 
were present, increasing the sulfide con- 
centration in the solution increased mer- 
cury extraction. This was probably due 
to the formation of a soluble HgS2^" com- 
plex (6^) . Because of this complex forma- 
tion, high concentrations of Na2S have 
been used to extract mercury from concen- 
trates ij) , However, care must be taken 
to control the amount of sulfide added, 
since the addition of too much Na2S may 
increase mercury extraction instead of 
precipitating mercury as desired. 

To determine the effects of Na2S con- 
centration and time on mercury precipita- 
tion (as HgS), additional tests were con- 
ducted. Solutions containing 0.34 Ib/st 
NaCN were contacted with gold ore for 24 
h to extract 12 pet of the mercury; then 
Na2S was added. Using ore containing on- 
ly 0.02 Ib/st Na2S, 79.9 pet of the sol- 
ubilized mercury was precipitated. How- 
ever, within 0.5 h, the precipitated mer- 
cury began to redissolve, and 4 h later, 
nearly 30 pet of it had redissolved. 

Because the HgS redissolved, CaS was 
investigated as an alternative. In one 
test, a cyanide solution containing 0.34 
Ib/st NaCN (with no sulfides) extracted 
12 pet of the mercury in 24 h. After 
adding only 0.02 Ib/st CaS to this solu- 
tion, 99.8 pet of the mercury was precip- 
itated in only 0.5 h, and 7 h later, only 

7 pet of the precipitated HgS had redis- 
solved. A followup test, using 0.09 lb/ 
St CaS, precipitated 100 pet of the solu- 
ble mercury; after 24 h, < 0.01 pet of 
the HgS had redissolved. 

A comparison between HgS redissolution 
using Na2S and CaS is shown in figure 2. 



About 10 times more HgS redissolved with 
Na2S than with CaS after 4 h. This may 
be due to the formation of the soluble 
Na2HgS2 complex (6^). CaS is less likely 
to form this type of complex because of 
its insolubility. 




0.02 



0.08 



0.10 



0.04 0.06 
Na2S, Ib/st ore 

FIGURE 1. - Effect of NojS and NaCN on mercury 
extraction. 



60 


1 


1 1 


1 1 


1 


u 






^ 




a. 




^^ 






q" 




^y^ 






UJ 




/^NogS 






rl 20 


- 






— 






y^ 






in 




/ 






Q 










UJ 










cr 










^ 10 


- X 






— 


z> 










u 










cc 

LlI 






^_CaS_______ 


— — " 


2 


' \ ' 


1 1 


1 i 


, 



3 4 

TIME, h 



FIGURE 2. - Effect of time on sulfide precipita- 
tion of mercury. 



21 



MERCURY PRECIPITATION AND CARBON ADSORPTION 



SIMULATED CARBON-IN-PULP 



LABORATORY GRINDING CIRCUIT 



Because most gold operations use acti- 
vated carbon to recover precious metals, 
tests were conducted to determine the ef- 
fect of sulfide addition on carbon ad- 
sorption of gold and mercury In a labora- 
tory carbon-ln-pulp (CIP) circuit. Five 
hundred grams of dry ground gold ore was 
leached with 870 mL of leach solution 
containing 0.34 ib/st NaCN and enough 
lime to give a pH of 11. The slurry was 
rolled for 24 h, then various amounts of 
CaS were added and rolling was continued 
for 1 h. Following HgS precipitation, 
18 g/L activated carbon was added and 
mixed for 1 h. The carbon was screened 
from the slurry, and the slurry was con- 
tacted with fresh carbon a second and 
third time. 

Results of these tests are listed in 
table 1, Gold and silver adsorption on 
carbon without prior CaS addition was 
nearly 100 pet after the first stage, 
while only 61 pet of the soluble mercury 
adsorbed. Additional adsorption stages 
recovered 9 pet more of the mercury. 
With 0.023 ib/st CaS, mercury adsorption 
was reduced to about 5 pet, without 
decreasing gold and silver adsorption; 
however, with 0.047 Ib/st CaS addition, 
there was a slight decrease in silver ad- 
sorption and mercury adsorption decreased 
to 0.9 pet. 

TABLE 1. - Carbon-ln-pulp adsorption 
with CaS precipitation, percent' 



Carbon stage 


Au 


Ag 


Hg 


Ib/st CaS: 

1 


99 
1 



99 
1 



99 
.6 
.1 


97 
3 


96 

4 


95 
4.1 
.5 


61 


2 


7 


3 


2 


0.023 Ib/st CaS: 

1 


4.7 


2 


.4 


3 


.3 


0.047 Ib/st CaS: 

1 


.37 


2 


.31 


3 


.20 



Several gold operations add NaCN and 
lime to the grinding circuit to extract 
gold and silver. Because some of the 
mercury is also extracted at the same 
time, tests were conducted to determine 
if addition of CaS to the ball mill would 
prevent mercury extraction and how it 
would affect gold, silver, and mercury 
loading on carbon. The following were 
added to a laboratory ball mill and 
ground for 45 mln: 1,000 g of minus 10- 
mesh ore, 0.5 Ib/st NaCN, 1,000 mL H2O, 
enough lime to give a pH of 11, and vary- 
ing amounts of CaS. Each resulting slur- 
ry was washed into a 9-L bottle with 
2,000 mL H2O and mixed. Gold and silver 
were removed from the slurry by adsorp- 
tion on activated carbon in three stages. 
One gram of fresh carbon was used in each 
stage. The total carbon contact time was 
24 h. 

The results of these tests are given in 
table 2. Mercury extraction and adsorp- 
tion on carbon were decreased from 15 to 
0.4 pet and from 5.5 to 0.17 pet, respec- 
tively, by increasing the amount of CaS 
from 0.012 to 0.096 ib/st. Gold extrac- 
tion decreased slightly, but gold loading 
on carbon was essentially unaffected. 
Silver extraction was unaffected, but 
silver adsorption decreased. Both gold 
and silver adsorption on carbon was lower 
than in previous sulfide precipitation 
tests. This was because much less car- 
bon was used (0.33 g/L) than is normal- 
ly used (18 g/L). Decreasing the carbon 

TABLE 2. - Effect of CaS addition during 
grinding on mercury, gold, and silver 
extraction and adsorption, percent 





Extraction' 


Ad 


sorption 


CaS, Ib/st 


Au 


Ag 


Hg 


on carbon^ 




Au 


Ag 


Hg 


0.012... 


99 


86 


15 


91 


71 


5.5 


.024 


94 


98 


1.5 


87 


74 


.55 


.048 


94 


98 


2 


88 


67 


.40 


.096 


95 


98 


1.4 


89 


64 


.17 



'Percent of soluble metal (prior to CaS 
addition) adsorbed on carbon. 



'Total extraction including grinding 
(45 mln) and carbon addition (24 h) . 
^24 h contact with carbon. 



22 



concentration made it possible to analyze 
for the small amount of mercury adsorbed 
on the carbon. 

GOLD- SILVER ORES 

Because silver was precipitated dur- 
ing mercury precipitation with CaS, addi- 
tional tests were conducted. The pur- 
pose of these tests was to determine the 
effects of adding CaS and to find a way 
to prevent silver precipitation. The 
ores used in this study are listed in 
table 3. 

TABLE 3. - Analyses of gold-silver ores 



Ore 


Hg, 
ppm 


Ag, 
tr oz/st 


Au, 
tr oz/st 


Cu, 
pet 


Carline. . . . 

Cortez 

Silver Reef 


16 
9 
3 


2.52 
.17 
10.1 


0.098 
.013 
.01 


NAp 
NAp 
0.8 



NAp Not applicable. 

The ores were leached with 1 Ib/st NaCN 
for 24 h; then CaS was added and mixed 
in for 1 h. Following mercury precipi- 
tation, the slurry was contacted with 
activated carbon for 1 h. The test re- 
sults are listed in table 4. Without 
CaS addition, gold, silver, and mercury 
adsorption values were all 100 pet; how- 
ever, with CaS addition, both the silver 
and mercury adsorption values were great- 
ly reduced for all ores except the Silver 
Reef ore. 

Tests were conducted to determine why 
silver was not precipitated from the 
Silver Reef slurry during mercury pre- 
cipitation. Close analysis of the Sil- 
ver Reef pregnant leach solution showed 
that it contained 170 ppm Cu. To elim- 
inate silver loss during mercury pre- 
cipitation, various amounts of CuCN 
were added to a pregnant leach solution 
containing 1,5 ppm Au, 3 ppm Ag, and 
0.76 ppm Hg prior to the addition of 0.1 
Ib/st CaS. The results given in figure 3 
show that the addition of 160 ppm Cu 



eliminated silver loss; but the copper 
addition did not affect precipitation of 
the mercury. 

Additional tests were conducted as 
described in table 4, except that 235 ppm 
Cu was added prior to mercury precip- 
itation. The results listed in table 5 
show that silver recovery was greatly 
increased. Carlin and Cortez silver 
adsorption was increased from to 80 pet 
and 6 to 91 pet, respectively. Mercury 
adsorption was zero for all ores. 

TABLE 4, - Adsorption on carbon in 
simulated three-stage carbon-in-pulp 
operation with CaS precipitation, 
percent' 



Carbon stage 



Au Ag Hg 



CARLIN ORE 



Ib/st CaS: 

1 


90 
9 

1 

50 

41 

9 


91 
9 







100 


2 





3 





0.10 Ib/st CaS: 

1 


8.1 


2 





3 






CORTEZ ORE 



Ib/st CaS: 

1 


57 
30 
13 

64 
25 
11 


75 

25 



6 




100 


2 





3 





0.10 Ib/st CaS: 

1 


2 


2 





3 






SILVER REEF ORE 






Ib/st CaS: 

1 


100 



100 




99 

1 


51 
25 
14 


100 


2 





3 





0.10 Ib/st CaS: 

1 





2 





3 






' Percent of soluble metal adsorbed on 
carbon. 



23 




50 



200 



75 100 125 150 175 
COPPER ADDITION, ppm 
FIGURE 3. - Effect of copper addition on silver 
precipitation during CaS precipitation of mercury. 



TABLE 5. - Effect of copper addi- 
tion on adsorption on carbon, 
percent^ 

(CaS addition: 0.1 Ib/st) 



Carbon stage 



Carlin ore 



Au Ag Hg 



Cortez ore 



Au Ag Hg 



WITHOUT COPPER ADDITION 



1 


50 

41 

9 







8.1 






64 
25 
11 


6 





? 


2 





3 










WITH 235 mg/L COPPER ADDITION 



1 


93 
13 

7 


60 
13 

7 







87.5 
.8 
.8 


75 
8 
8 





2 





3 










^Percent of soluble metal adsorbed on 
carbon. 



MERCURY PRECIPITATION IN A MILL OPERATION 



Precipitation of mercury with sulfides 
is being tested at a northern Nevada mill 
operation, using a process similar to 
that illustrated in figure 4. Because 
CaS was difficult to obtain, sodium hy- 
drogen sulfide (NaHS) is being used in- 
stead. A solution containing NaHS is 
added to the NaCN slurry from the ball 
mill before the slurry enters the thick- 
ener. As shown in table 6 (feed to 
thickener, days 3 to 6) , nearly all of 



the mercury is precipitated at the point 
of addition; however, over time, about 30 
to 50 pet of this precipitated mercury 
(HgS) redissolved. This was not unex- 
pected, as NaHS acts similarly to Na2S, 
and more NaHS addition points are re- 
quired to obtain a more complete mercury 
precipitation. Possible addition points 
include the final leach tank prior to the 
CIP tanks and the CIP tanks. 



TABLE 6. - Mercury precipitation with NaHS in a mill operation 
(Mercury analyses, parts per million^) 



Sampling area 


Sampling day 




1 


2 


3 


4 


5 


6 


Feed to thickener 


2.5 


2,6 


0.08 


0,02 


0.05 


0.02 


Feed to carbon column 














(thickener overflow) 


2.6 


2.7 


.09 


.07 


.11 


.01 


Final carbon column. ........ 


2.6 


2.8 


.16 


.03 


.09 


,02 


Feed to leach tanks (thick- 




ener underflow) 


2.5 


2.8 


,68 


.47 


.83 


,19 


Feed to carbon in pulp (from 














leach tank 4) 


2.7 


2.9 


3,3 


1.3 


1.8 


1,3 


Final carbon in pulp (to 














tailings) 


2,2 


2.8 


3,2 


1.5 


1.4 


1,2 



'Mercury remaining in solution. 



24 



Ore »■ 



NaHS 




Carbon regeneration 
kiln 



Tailings 
(HgS) 



Return carbon 



Mercury retort 



Smelting furnace 



Mercury 



FIGURE 4. 
operation. 



Precious metals 
- Flowsheet showing NaHS addition points for mercury precipitation in a mil 



CONCLUSIONS 



25 



CaS addition to gold-silver cyanide 
leach slurries or to a laboratory grind- 
ing circuit containing NaCN and lime re- 
duced mercury dissolution to < 0,5 pet. 
Mercury loading on activated carbon was 
reduced to < 0.2 pet. Gold loading on 
activated carbon was affected very little 
by the sulfide addition; however, silver 
loading appeared to be affected. Addi- 
tional NaCN-CaS CIP tests with ores con- 
taining greater amounts of silver result- 
ed in only to 6 pet of the silver being 



adsorbed by the carbon rather than the 
normal 90 to 100 pet. Addition of copper 
as CuCN, prior to the CaS addition, re- 
sulted in 80 to 90 pet of the silver and 
none of the mercury being adsorbed on the 
carbon. 

Preliminary testing in a mill operation 
using NaHS showed mercury precipitation 
was nearly complete at the point of addi- 
tion; however, as had been found in other 
tests with Na2S, 30 to 50 pet of the pre- 
cipitated HgS redissolved with time. 



REFERENCES 



1. Craft, W. B. The Pinson Gold 
Story. AZ Conf., AIME, Tucson, AZ, Dec. 
6, 1982, 8 pp.; available upon request 
from R. G. Sandberg, BuMines, Salt Lake 
City, UT. 

2. Skillings, D. N. , Jr. Getty Mining 
Co. Starting Up Its Mercury Gold Opera- 
tion in Utah. Skillings' Min. Rev., v. 
72, No. 17, 1983, pp. 4-9. 

3. . Homes take Proceeding With 

Its McLaughlin Gold Project. Skillings' 
Min. -Rev., v. 72, No. 4, 1983, pp. 3-6. 

4. Sandberg, R. G. , W. W. Simpson, 
and W. L. Staker. Calcium Sulfide 



Precipitation of Mercury During Cyanide 
Leaching of Gold Ores. BuMines RI 8907, 
1984, 13 pp. 

5. Flynn, C. M., Jr., T. G. Carnahan, 
and R. E. Lindstrom. Selective Removal 
of Mercury From Cyanide Solutions. U.S. 
Pat. 4,256,707, Mar. 17, 1981. 

6. Habashi, F. Hydrometallurgy . 
Principles of Extractive Metallurgy, v. 
2. Gordon and Breach, 1969, p. 99. 

7. Town, J. W. , and W. A. Stickney. 
Cost Estimates and Optimum Conditions for 
Continuous-Circuit Leaching of Mercury, 
BuMines RI 6459, 1964, 28 pp. 



26 



CARBONACEOUS GOLD ORES'" 
By B, J. Schelner2 



ABSTRACT 



The presence of organic material in 
gold ores interferes with gold extrac- 
tion by cyanidation. Oxidation of the 
organic material using chlorine-hypo- 
chlorite systems has been shown to render 
the gold ore amendable to cyanidation. 
This paper discusses three oxidation 
procedures that have been investigated 
by the Bureau of Mines: (1) addition of 



sodium hypochlorite (NaOCl) to ore pulp, 

(2) addition of chlorine to ore pulp, and 

(3) in situ generation of NaOCl by elec- 
trolysis of a brine solution used to pulp 
the ore. Gold extraction was greater 
than 90 pet when carbonaceous ore was 
oxidized and then cyanided. Both labo- 
ratory and pilot plant results are 
discussed. 



INTRODUCTION 



The presence of carbon and organic com- 
pounds that inhibit gold recovery from 
auriferous ores has long plagued cyanide 
mill operators. Research to develop 
techniques for testing these ores, which 
are commonly referred to as carbonaceous 
ores, was conducted in the early 1920' s 
by the Bureau of Mines (O.-^ Research 
was continued by the Bureau and others, 
but as late as 1958, it was indicated 
that no solution to the carbonaceous 
problem was at hand (2^). After the dis- 
covery of carbonaceous gold ores at the 
Carlin Mine in northeastern Nevada, the 
Bureau initiated an extensive research 
program in 1966 to solve the problems as- 
sociated with extracting gold from these 
ores. 

Research was conducted by the U.S. 
Geological Survey, the Bureau of Mines, 
and Newmont Mining Corp., the owners of 
the Carlin Gold Mine, to characterize 
the carbonaceous ores found in Nevada 
(3-5). Early studies indicated that the 



gold-bearing sedimentary beds were asso- 
ciated with the Roberts Mountain thrust 
fault system, generally in windows expos- 
ing lower plate sedimentary beds of Silu- 
rian age. However, later discoveries 
have shown that the ore association with 
the Roberts Mountain thrust fault system 
is just a fortuitous anomaly, since gold 
has been found in upper plate material 
outside the Roberts Mountain thrust 
fault. It has been hypothesized that the 
gold was redistributed and concentrated 
in permeable carbonaceous horizons by 
hydrothermal solutions. Apparently the 
same solutions replaced much of the silty 
dolomitic limestone with microcrystalline 
quartz. Subsequent oxidation induced by 
shallow meteoric oxygen-bearing waters 
removed the carbonaceous material from 
the upper portions of the deposits, thus 
oxidizing these portions of the deposits. 
The Carlin Gold Mine property is an exam- 
ple of a deposit that contains both oxi- 
dized and carbonaceous gold ores. 



CYANIDATION OF CARBONACEOUS ORES 



To determine the effect of carbonace- 
ous ores on cyanidation, samples of both 
oxidized and carbonaceous ore were ob- 
tained from various locations in Nevada, 
including the Carlin Gold Mine, Cortez 

^This paper is based upon work done un- 
der an agreement between the University 
of Alabama and the Bureau of Mines, 

^Supervisory metallurgist, Tuscaloosa 
Research Center, Bureau of Mines, Univer- 
sity, AL. 



Gold Mine, and the abandoned Gold Acres 
pit at Crescent Valley. The gold content 
of the carbonaceous ores ranged from 0.23 
to 0.40 tr oz/st Au, with an organic 
carbon content ranging from 0.3 to 0,6 
wt pet. 

Initial experiments involving cyanida- 
tion of the carbonaceous ores over a wide 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



27 




10 20 30 40 50 

CARBONACEOUS ORE ADDED TO OXIDE ORE, wt pet 

FIGURE 1. - Effect on gold extraction of adding 
carbonaceous ore to oxide ore. 

range of operating conditions showed that 
only 5 to 32 pet of the gold was amenable 
to recovery by this conventional tech- 
nique. The data indicated that cyanide 
was being consumed, but not excessively. 
The oxygen content of the leach solution 
was determined to be favorable for effec- 
tive cyanidation. Analysis of washed 
tails detected cyanide, suggesting that 
Au(CN)2- was being adsorbed on carbonace- 
ous components of the ore. The possibil- 
ity of gold adsorption by carbon was fur- 
ther investigated by contacting 0.1 tr/st 
of pregnant Au(CN)2- solutions with vari- 
ous carbonaceous ores. From 12.5 to 140 
tr oz Au was adsorbed per short ton of 
carbonaceous ore. 

To investigate how the addition of 
carbonaceous material affects the cya- 
nidation of oxidized ores, a series of 
experiments was conducted in which 
various amounts of carbonaceous material 
were added to oxidized ore and the sam- 
ple cyanided. The results of these 
experiments are shown in figure 1. Gold 
extraction decreased linearly as the car- 
bonaceous ore content increased (_6-_7) . 
Application of the well-known technique 
of blanking out an activated carbon com- 
ponent by coating it with kerosene or 
some other petroleum product was shown 
to be only partially effective, depend- 
ing on the particular carbonaceous ore 
being treated. Treatment of the ore 
with anion-exchange resins or granulated 



activated carbon during cyanidation 
in order to actively compete for the 
Au(CN)2- complex generally improved the 
gold extraction, but the results varied 
considerably depending on the particular 
sample being used. This suggested that 
certain samples of the carbonaceous mate- 
rial contained two different types of 
carbon that prevented favorable extrac- 
tion: (1) an activated type of carbon 
that physically adsorbed the Au(CN)2- 
complex, and (2) a hydrocarbon type that 
formed a gold compound during deposition 
that was not attacked by cyanide (^-7^) • 

Gold extraction depended markedly on 
the basicity of the system, indicating 
the presence of carboxylic acid groups 
that saponified in the highly basic ion- 
exchange system. Isolation of organic 
compounds in the sample was accomplished 
by an aqueous NaOH treatment for 2 h, 
followed by acidification of the liquor 
and solvent extraction of the dissolved 
organic compounds into chloroform (3) . 
Infrared spectra of the organic extrac- 
tion product had major adsorption peaks 
at 2,900 and 1,700 cm~^ , which were char- 
acteristic of long-chain carboxylic 
acids. The neutralization equivalent of 
this mixture of organic acids was 1,500. 
Sulfur and nitrogen groups were also 
found in the organic material. The 
amounts of organic material that could 
be accumulated from the sample were only 
sufficient for identification purposes. 
The extracted organic compounds were 
found to be remarkably similar to humic 
acid extracted from leonardite that oc- 
curs in North Dakota (8^). These humic 
acid extracts contain long-chain carbox- 
ylic acids as well as sulfur and nitrogen 
groups. Cyanidation of oxidized ore in 
the presence of humic acids showed that 
Au(CN)2- was adsorbed or associated with 
the humus compounds and that gold recov- 
ery was essentially nil. 

It was concluded that the low extrac- 
tion of gold from the carbonaceous mate- 
rials could not be attributed wholly to 
physical adsorption of the Au(CN)2- com- 
plex on carbon, but that in certain of 
these materials, a substantial amount of 
the gold was locked in in the form of a 
chelate containing CO-N-S ligands. 



28 



CHEMICAL OXIDATION OF CARBONACEOUS ORES 



LABORATORY EXPERIMENTS 

The data indicated that destruction or 
complete pacification of detrimental car- 
bon components is a requisite for effec- 
tively extracting gold from the various 
carbonaceous materials. It was deter- 
mined that a mild chemical oxidation of 
the carbonaceous ores followed by cya- 
nidation yielded high gold extraction. 
Various oxidation systems were investi- 
gated, including ozone, chlorine, NaOCl, 
calcium hypochlorite [Ca(0Cl)2], perman- 
ganates, perchlorates, chlorates, and 
oxygen. Gold extraction, obtained by 
cyanidation after oxidation, increased 
significantly in nearly every case over 
that obtained without pretreatment ( 6-_7 ) . 

The use of ozone as an oxidant was fur- 
ther investigated (9^). It was determined 
that 95 pet Au extraction could be ob- 
tained from several different carbonace- 
ous ores by slurrying the ore with brine 
solution, lowering the pH to 1 with H2S0i^ 
or HNO3 , and bubbling ozone through the 
slurry. After the ozone treatment was 
completed, the gold was extracted by cya- 
nidation. However, owing to the acid- 
consuming calcareous nature of the ore, 
the ozone system, in which an acid medium 
was used, was ruled out as an impractical 
solution to the carbonaceous problem. 

Introduction of chlorine gas into an 
ore slurry, followed by filtering and 
subsequent cyanidation of the filter 
cake, resulted in 95 pet Au extraction 
from several carbonaceous ores (_7 ) . 
Rapid addition of the chlorine gas (as 
represented below) caused the pH of the 
ore pulp to drop to the range of pH 1 
to 2. 



CI2 + H2O ->■ HCl + HOCl. 



(A) 



However, the pH remained neutral with 
slow addition of chlorine gas because the 
HCl formed reacted slowly with the cal- 
careous gangue material to give CaCl2 : 

2HC1 + CaC03 ->■ CaClg + H2CO3. (B) 

The active oxidizing species obtained 
from rapid addition of chlorine gas 
is principally HCl. Slow addition of 



chlorine gas results in production of 
Ca(0Cl)2 as the principal oxidizing 
species: 

2H0C1 + CaC03 -»■ Ca(0Cl)2 + H2CO3. (C) 

The amount of chlorine gas that escapes 
from the pulp can be excessive. To over- 
come this problem, NaOH or lime was added 
to the slurry before the addition of 
chlorine gas. The chlorine gas reacts 
with the base to form Ca(0Cl)2 or NaOCl: 

CI2 + 2NaOH -»- NaOCl + NaCl + H2O, (D) 

2CI2 + 2Ca(0H)2 ->■ Ca(0Cl)2 

+ CaCl2 + 2H2O. (E) 

The hypochlorite or oxidizing species 
then reacts with carbonaceous material in 
the ore, passivating activated-type car- 
bon and breaking hvdrocarbon chains and 
htunic acid-type components. 

Since the NaOCl (household bleach) 
proved to be an effective oxidizing 
agent, the parameters affecting gold ex- 
traction were investigated. It was de- 
termined that 90 pet Au extraction could 
be obtained at 50° C in 4 h using 16 to 
20 lb NaOCl per short ton of ore along 
with 20 Ib/st of lime (7^). 

Instead of adding NaOCl or chlorine to 
the ore pulp, the oxidizing condition 
could be produced by slurrying the ore in 
brine and electrolyzing to produce NaOCl 
in situ. A literature survey indicated 
that the technology for the electrolysis 
of NaCl in ore slurries was essentially 
nonexistent. Preliminary experiments in- 
dicated that oxidizing conditions could 
be generated and controlled under a vari- 
ety of conditions by this electrooxida- 
tion concept (10-11). Parameters deemed 
to be important to the development of the 
technique included salt concentration, 
electrolysis time at constant amperage 
per short ton of ore, temperature, cur- 
rent density, type of electrodes, elec- 
trode spacing, and particle size of ore 

(I). 

During the course of the investigation 
it was determined that a variety of elec- 
trode materials and configurations could 



29 



be used. The only difficulty encountered 
was a buildup of a deposit on the cathode 
with time. This difficulty was overcome 
by using graphite electrodes and revers- 
ing the current periodically to remove 
the deposits. One type of cell used, a 
plate-type laboratory-scale electrooxida- 
tion cell, is shown in figure 2. 

The effect of salt concentration (NaCl) 
on gold extraction from carbonaceous ore 
was investigated. It was determined that 
an 8- to 10-pct-salt concentration was 
adequate for obtaining sufficient oxida- 
tion to result in efficient gold extrac- 
tion by cyanidation. 

The effect of temperature on gold ex- 
traction was determined through tests at 
30°, 40°, and 50° C. Maximum gold ex- 
traction was obtained at 40° C. Heat in- 
put to an electrooxidation system and the 
resulting temperature are functions of 
the conductivity of the electrolyte and 
the power required to accomplish the oxi- 
dation; therefore, parameters such as 
electrode spacing, salt concentration, 
and pulp density are all critical in 
maintaining the desired temperature. 
Generally speaking, electrode spacing 
should be as close as is consistent with 



Busbar 



Pulp level 



Wood 
support 



Graphite 
cathode 

Graphite 
anode 



Ore slurry 

30-50 pet 

solids 




good pulp flow between electrodes. The 
effect of increasing electrode spacing 
was investigated. Spacings of 3/8, 5/8, 
and 1-1/8 in were used, with pulp den- 
sity at 40 pet and a salt concentration 
of 10 pet. As was expected, the resist- 
ance between the electrodes increased 
with increased electrode spacing. The 
effect of pulp density on conductance was 
also studied (fig. 3). As was expected, 
the conductance decreased rapidly as pulp 
density increased. 

Current density is another factor 
that affects the voltage-amperage 
relationship, which in turn affects the 
temperature. The effect of current den- 
sity on the voltage required to maintain 
constant amperage during electrolysis is 
shown in figure 4. The voltage increased 
linearly with current density over the 
range measured. The data indicated that 
the current density should be as low as 
possible to keep power consumption at a 
minimum. However, the current density 
dictates the number of electrodes re- 
quired, and the tonnage of the mill and 
the size of the agitators employed are 
important factors that must also be con- 
sidered in projecting the number of elec- 
trodes that can be practically utilized. 

Grinding is an important part of any 
hydrometallurgical process; it releases 



0.18 




FIGURE 2. - Plate-type laboratory-scale electro- 
oxidation cell and agitation vessel. 



20 30 40 50 60 

PULP DENSITY, pot 

FIGURE 3. - Effect of pulp density on conductance 
in ore-brine pulp (using 9.77-pct-salt solution). 



30 




0.4 0.8 1.2 1.6 2.0 

ANODE CURRENT DENSITY, A/in^ 

FIGURE 4. - Effect of anode current density on 
voltage at constant amperage. 

the mineral from the host rock so the 
mineral can come in contact with the re- 
actants. The effect of particle size on 
gold extraction was investigated using 
several different grinds based on the 
percentage of minus 200-mesh material. 
Data for one ore indicated that gold ex- 
traction increases as the particle size 
becomes smaller and that a grind of 70 
pet minus 200 mesh was satisfactory; how- 
ever, each ore must be evaluated on an 
individual basis as to optimum particle 
size for reaction. 

PILOT PLANT STUDIES 

Based on the results obtained in labo- 
ratory studies, the use of NaOCl and 
chlorine and the in situ generation of 
NaOCl (electrooxidation) were investi- 
gated on a pilot plant scale (12-13). 
The ores used in these studies were ob- 
tained from the Carlin Gold Mine. The 
flow sequence and pilot plant operation 
generally followed conventional counter- 
current-decantation slime-circuit operat- 
ing practice. Ore material was ground to 
60 pet minus 200 mesh at 50 pet solids in 
the rod mill and pulped into the oxida- 
tion tanks, each of which held 275 lb of 
dry ore (550 lb of pulp). After treat- 
ment in the oxidation tank at the desired 
temperature, the pulp was passed through 
the digestion-surge tank, where 1 lb of 
cyanide per short ton of ore was added, 
along with sufficient lime to maintain a 
cyanidation pH of 11. Approximately 5 lb 
of lime per short ton of ore was usually 



required to maintain the desired pH val- 
ue. The pulp was then pumped through 
three cyanidation tanks, for a cyanida- 
tion time of 9 h. From the cyanidation 
tanks the pulp passed through four thick- 
eners at 20 pet solids, flowing counter- 
current to the barren solution recycled 
from the gold-precipitation sequence. 
Pregnant solution was passed from the 
first thickener to the gold-precipitation 
system. 

Initial pilot plant experiments were 
conducted on nonrefractory oxide gold ore 
to determine the best operating condi- 
tions for the grinding circuit and pulp 
handling, optimum flow rates, etc. Mill 
tails from the oxide ore contained 0.008 
tr oz/st Au, which corresponds to a gold 
extraction of 96 pet. 

Carbonaceous gold ore was processed in 
the pilot mill, using conventional cyani- 
dation, without oxidation pretreatment , 
as a baseline experiment to determine the 
effect of carbonaceous material. Gold 
extraction obtained in these experiments 
ranged from 29 to 33 pet (0.26 to 0.28 
tr oz Au per short ton of tails). 

Oxidation by NaOCl Addition 

The pilot plant operations generally 
followed the operating conditions estab- 
lished in the laboratory experiments for 
obtaining favorable gold extraction val- 
ues. The oxidation section of the plant 
was operated on a semicontinuous basis 
with ground ore being treated on a batch 
basis in successive oxidation tanks. The 
ore then could be discharged into the 
surge tank in such a manner that a re- 
serve of treated pulp was available for 
continuous operation of the cyanidation 
and liquid-solids separation sections of 
the mill. The oxidation section of the 
mill processed 80 lb of dry ore per hour, 
thus providing a retention time of 9 h in 
the cyanidation tanks. 

Figure 5 shows the oxidation results — 
the gold extractions obtained at differ- 
ent levels of NaOCl addition to the pulp 
containing carbonaceous ore. As shown in 
figure 5, the reaction time, temperature, 
and lime addition were kept constant. 
The pulp pH values were in the 11 to 11.5 
range. Extraction increased markedly 



31 



90 



' ' ^— — 


- 


_ / Conditions: 




/ Reaction time, 4 h 




/ Lime, 10 Ib/st 




/ Temp, 50° to 60° C 




/ill 


— 



^ 80 
o 

a. 

o 70 

I- 

^ 60 

X 

UJ 

Q 50 

_i 
o 
o 
40 



30. 

5 10 15 20 

NoOCI ADDITION, Ib/st 

FIGURE 5. - Effect of NaOCI addition on gold 
extraction. 

with addition of 5 to 10 Ib/st NaOCl and 
then increased gradually with larger 
NaOCl additions. Gold extraction by sub- 
sequent cyanidation reached 90 pet with 
a 20-lb/st-NaOCl addition. The results 
closely paralleled similar laboratory ex- 
periments conducted on a different sample 
of carbonaceous gold ore. 

In other experiments, a reaction time 
of 3 h (instead of 4 h) was shown to be 
sufficient when 20 lb of lime per short 
ton was used. It was also determined 
that the optimum temperature for oxida- 
tion with NaOCl lies between 50° and 
60° C. 

Oxidation by Chlorine Addition 

Addition of chlorine to a pulp of fine- 
ly ground carbonaceous ore and water was 
investigated as a means of producing hy- 
pochlorite oxidant in situ on an economi- 
cal and easily controlled basis. As 
chlorine is bubbled into the pulp, it re- 
acts with water to form HOCl and 
HCl. These products are buffered by the 
calcareous gangue in the ore to form 
Ca(0Cl)2 and CaCl2 . Oxidation of the 
carbonaceous matter is thought to be ac- 
complished by the hypochlorite ion. 

It had been shown in the laboratory ex- 
periments that addition of chlorine gas 
at moderate rates to the agitated pulp 
through a sintered-disk sparger resulted 
in favorable oxidation without excessive 



loss of chlorine gas. The chlorine re- 
acted rapidly, and gold extraction from 
the oxidized pulp by subsequent cyanida- 
tion was shown to be largely independent 
of the rate of chlorine addition. The 
factor limiting the rate of chlorine ad- 
dition appeared to be the amount of hypo- 
chlorite remaining after oxidation. The 
hypochlorite product reacted with the 
carbonaceous matter as chlorine was added 
during the initial stages of the oxida- 
tion, and no hypochlorite could be de- 
tected in the solution. The amount of 
hypochlorite in solution increased with 
time until the oxidation was completed. 
It was determined that the oxidation can 
be completed in as little as 4 h; how- 
ever, at this relatively rapid rate of 
chlorination, the Ca(0Cl)2 in solution 
can increase to 1 pet or more in the 
later stages of oxidation. If hypochlo- 
rite is allowed to rise to a level that 
cannot be consumed by the ore, it will 
constjme cyanide in the cyanide circuit. 
Therefore, it is desirable to limit the 
Ca(0Cl)2 in solution to less than 0.2 pet 
during chlorination. This resulted in a 
chlorination time of up to 15 h at 24° to 
30° C for some of the more refractory 
samples, with an additional 5 h for the 
ore to consume the residual hypochlorite. 
The final concentration of Ca(0Cl)2 going 
to cyanidation was less than 0.02 pet. 

Pilot plant tests were run with three 
55-gal drums connected in series, each 
containing 275 lb of solids at 40 pet 
solids. Chlorine was fed to each drum 
through a 1/2-in-diam pipe extending to 
the bottom of the drum. Oxidation was 
conducted on a batch basis with the pulp 
overflow from the third tank being con- 
tinuously recirculated through the system 
to give the desired chlorination time. 
The near-optimum temperature of 24° to 
30° C was maintained by heating the pulp. 
The pH of the system remained at 6,6 dur- 
ing chlorination without adjustment. 
Figure 6 shows that gold extraction in- 
creased rapidly with increasing chlorine 
addition. The optimum extraction 92 pet, 
was obtained using 40 Ib/st chlorine. 

Scale-up experiments were conducted in 
a 6- by 7-ft tank containing 7.5 st of 
pulp at 40 pet solids. Chlorine gas was 
bubbled in through four pipes submerged 



32 



t3 100 



80- 





1 1 

^Q^ — ' ° 




Y 




^^"-"-''^ Conditions: 








Reaction time, 


15 h 






pH, 6.6 








Temp, 24° to 3C 

1 1 


"C 





Cathode 
busbar 



20 



30 40 

CHLORINE, Ib/st 



50 



FIGURE 6. - Effect of chlorine addition on gold 
extraction. 

5 ft into the tank. Chlorine was added 
over an 8-h period at the rate of 3.5 lb/ 
(sfh) initially, then decreased to 1.5 
lb/(sfh) over the last 2.5 h. Gold 
extractions of 89 pet (0.035 oz Au per 
short ton of tails) were obtained using 
38 Ib/st chlorine. These results paral- 
leled those obtained in the pilot-plant- 
scale experiments. 

Electrooxidation of Carbonaceous Ores 

In additional pilot plant studies, 
electrolysis used on ore pulp prepared 
from finely ground carbonaceous ore and 
salt brine was investigated as a means of 
generating oxidizing conditions in situ. 
The cell used was a simple plate-type 
arrangement consisting of graphite elec- 
trodes 2-1/2 in wide, 3/4 in thick, and 
30 in long (fig. 7). Identical graphite 
cathodes and anodes were placed alter- 
nately in a nonconductive holder with 
1/2-in spacing, which allowed favorable 
pulp flow through the system. 

The pilot plant investigations deter- 
mined the effect of temperature on gold 
extractions, the effect of salt concen- 
tration on gold extractions and power re- 
quirements, and the effect of electrode 
spacing at various salt concentrations 
on power requirements. The data obtained 
were similar to the data obtained in pre- 
vious laboratory tests. Gold extractions 
of 90 pet were obtained using 10-pct-salt 
concentrations at 40° C, a current den- 
sity of 0.67 A/in^ , and an electrode 
spacing of 1/2 in. 

Scale-up experiments were conducted on 
a batch basis using a 6- by 7-ft agitator 
tank capable of holding 7.5 st of pulp at 
40 pet solids. Electrooxidation was ac- 
complished with two banks of graphite 



Pulp level 



Anode 
busbar 



Electrodes 



Ore slurry 

30-50 pet 

solids 




FIGURE 7. - Plate-type electrode assembly for di- 
rect electrolytic production of NaOCl in ore slurry. 



90 



o 

Q. 



g 80 

< 

I- 
^ 70 

Q 
_1 
O 
CD 



60 



1 


/ Conditions: 

2,800 A at 5.2 V 


=t.*- — 


1 


Temp, 40° C 

Salt cone, 10 pet 

1 1 1 





20 30 40 50 60 

ENERGY CONSUMPTION, kW-h/st 



70 



FIGURE 8. - Gold extraction and energy consump- 
tion for large-scale experiments. 

electrodes used in parallel, each con- 
taining 12 anodes and 11 cathodes. Tests 
were conducted at 40° C, with a 2,800-A 
current and a 10-pct-salt concentration 
in the pulp solution. Oxidized pulps 
were not cyanided in the pilot plant, but 
samples were taken at 1-h intervals and 
cyanided on a bench scale. Figure 8 
shows that gold extraction increased with 
electrolysis time, reaching a maximum ex- 
traction of 89.4 pet, which corresponded 
to 22 h of electrolysis. Data obtained 



33 



from the larger experiments indicated 
that scale-up to commercial plants should 
not present any serious problems. 

The pilot plant studies were a coopera- 
tive effort of the Newmont Mining Corp.'s 
Carlin Gold Mine and the Bureau of Mines. 
Experiments were conducted at the Carlin 



Gold Mine and at the Bureau's Reno (NV) 
Research Center. Based on the results of 
the testing program, the Carlin Mine in- 
stalled a full-scale facility using chlo- 
rine to treat carbonaceous ores prior to 
cyanidation. 



CONCLUSIONS 



Treatment of carbonaceous gold ores 
from north-central Nevada with NaOCl or 
chlorine, or by electrolytic oxidation 
resulted in favorable gold recovery by 
subsequent cyanidation. The gold that is 
complexed by humic acid- type compounds 
was liberated, and the adsorptive proper- 
ties of the ore were passivated by the 
oxidation treatment. Gold extractions in 
the 90-pct range were obtained on several 



tonnage ore samples from the Carlin Gold 
Mine. Equivalent metallurgical perfoirm- 
ance, compatible with present plant prac- 
tice, was obtained with NaOCl, chlorina- 
tion, and electrooxidation. The choice 
of one of these methods is therefore dic- 
tated by economic conditions such as the 
cost and availability of electric power, 
NaOCl, and chlorine. 



REFERENCES 



1, Leaver, E. S., and J. A. Woolf. 
Re-Treatment of Mother Lode (California) 
Carbonaceous Slime Tailings. BuMines 
Tech. Paper 481, 1930, 20 pp. 

2, Hedley, N, , and H, Tabachnick, 
Chemistry of Cyanidation, Mineral Dress- 
ing Notes, No. 23. Am. Cyanamid Co., New 
York, June 1958, 54 pp. 

3, Radtke, A, S,, and B, J, Scheiner, 
Studies of Hydrothermal Gold Deposition 
(I), Carlin Gold Deposit, Nevada: The 
Role of Carbonaceous Material in Gold 
Deposition, Econ, Geol, , v, 65, 1970, 
pp, 87-102, 

4, Radtke, A, S,, C. Heropoulos, B, P, 
Fabbi, B. J, Scheiner, and M, Essington. 
Data on Major and Minor Elements in Host 
Rocks and Ores, Carlin Gold Deposit, Ne- 
vada. Econ. Geol., v. 67, 1972, pp. 975- 
978. 

5, Hausen, D. M. , and P. F. Kerr. 
Fine Gold Occurrence at Carlin, Nevada, 
Sec, in Ore Deposits of the United 
States, 1933-1967, ed, by J, D, Ridge, 
AIME, V, 1, 1968, pp, 908-940, 

6, Scheiner, B, J. , R. E. Lindstrom, 
and T, A. Henrie. Investigation of Oxi- 
dation Systems for Improving Gold Recov- 
ery From Carbonaceous Materials. BuMines 
TPR 2, July 1968, 8 pp. 



and C. M. Frost, 
Byproduct, Bu- 

pp, 
R, E. Lindstrom, 



7, Scheiner, B, J,, R, E, Lindstrom, 
and T, A, Henrie, Oxidation Process 
for Improving Gold Recovery From Carbon- 
Bearing Gold Ores, BuMines RI 7573, 
1971, 14 pp, 

8, Fowkes, W, W, , 
Leonardite: A Lignite 
Mines RI 5611, 1960, 12 

9, Scheiner, B, J,, 
and T, A, Henrie, Process for Recovery 
of Gold From Carbonaceous Ores, U.S, 
Pat. 3,574,600, Apr. 13, 1971. 

10. Scheiner, B. J., R. E. Lindstrom, 
and T. A. Henrie. Electrolytic Oxidation 
of Carbonaceous Ores for Improving Gold 
Recovery. BuMines TPR 8, 1969, 12 pp. 

11. . Recovery of Gold From Car- 
bonaceous Ores. U.S. Pat. 3,639,925, 
Feb. 8, 1972. 

12. Scheiner, B, J,, R. E, Lind- 
strom, W. J. Guay, and D, G, Peterson, 
Extraction of Gold From Carbonaceous 
Ores: Pilot Plant Studies, BuMines RI 
7597, 1972, 20 pp, 

13. Scheiner, B, J,, R, E. Lindstrom, 
and T, A, Henrie. Processing Refractory 
Carbonaceous Ores for Gold Recovery. J. 
Met., V. 23, No. 3, 1971, pp. 37-40. 



34 



CARBON ADSORPTION-DESORPTION 
By J. A. Eisele"! 



ABSTRACT 



The Bureau of Mines has developed 
three methods for stripping gold and 
silver from activated carbon that enable 
the carbon to be reused many times. 
These three desorption techniques are to 
strip the precious metals with either 
(1) boiling NaOH-NaCN solution at atmos- 
pheric pressure; (2) NaOH-NaCN solution 



at temperatures above boling, in pressure 
vessels; or (3) an NaOH-NaCN solution 
containing ethyl alcohol at temperatures 
below boiling. All three methods are 
widely applied in industry. This paper 
describes the three methods and their 
development. 



INTRODUCTION 



The introduction of cyanide leaching in 
the 1890' s revolutionized the processing 
of gold and silver ores. Since that 
time, the standard hydrometallurgical 
process to recover gold and silver from 
disseminated ores has become cyanide 
leaching in a countercurrent decantation 
(CCD) plant (U.2 The operation of a 
conventional CCD plant (fig. 1) consists 
of the following steps: 

1. The ore is crushed and ground. 

2. During grinding, the ore is slur- 
ried with dilute basic cyanide solution 
typically containing 0.1 pet NaCN and 

'Supervisory chemical engineer, Reno 
Research Center, Bureau of Mines, Reno, 
NV. 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



c'ao" ^^°o'^' K-l Comminution I —\ Leaching lonk I I JLeoctiing lonk 2 



]^ 



t^ 



^7>^ 



Precipilalion \-^ Zinc du5 



Gold-silver sponge 



FIGURE 1. - Flow diagram of countercurrent decan- 
tation cyanide leaching plant. 



enough CaO to maintain a pH range of 10 
to 11. 

3. The slurry is agitated in a series 
of leaching tanks to give a total leach- 
ing time of 12 to 48 h, depending on the 
characteristics of the ore. 

4. The slurry is thickened and 
washed countercurrently in a series of 
thickeners. 

5. The pregnant solution from the 
first thickener is clarified by filtra- 
tion and deaerated by vacuum. 

6. Zinc powder is added to precipitate 
the gold and/or silver. 

7. The gold-silver-zinc sponge is fil- 
tered and refined to bullion, 

8. The barren cyanide solution is re- 
turned to the washing-leaching circuit, 
and makeup cyanide and base are added. 

Gold and silver recovery from pregnant 
solutions by zinc precipitation can pre- 
sent problems. Ideally, the pregnant so- 
lution should be clear before precipita- 
tion, and it, must be deaerated. This can 
be difficult to achieve with slimy ores. 
Also, precipitation is an inefficient 
method for recovering gold from dilute 
solutions. Different forms of carbon, 
especially activated carbon, have long 
been known to be good adsorbers of gold 
and silver cyanides from solution. How- 
ever, since the only known means to re- 
cover the gold and silver from the carbon 
was to burn it, carbon was not used ex- 
tensively. In a few cases, where slimy 
ores were being treated, gold adsorption 
on carbon was the preferred method for 



35 



recovering the precious metal values be- 
cause the clarification step was elimi- 
nated. The precious-metal-loaded carbon 



was recovered from the 
ing or flotation. 



slurry by screen- 



USE OF BOILING NaOH-NaCN SOLUTION 



In the late 1940' s, Bureau of Mines re- 
searchers were looking for ways to strip 
the precious metals from loaded carbon to 
enable the carbon to be reused a number 
of times, A World War II surplus of ac- 
tivated carbon manufactured from fruit 
pits was available at prices that made 
using carbon cheaper than using zinc, A 
means for desorbing and recycling the 
carbon would make carbon adsorption a 
very economical process for the recov- 
ery of gold. In 1950, an alkaline Na2S 
stripping method was described that 
eluted the gold, but not the silver, from 
carbon (2^), As most ores contain some 
silver, and silver in some ores is the 
primary value, the sulfide stripping 
method was not satisfactory because the 
carbon would eventually be loaded with 
unstrippable silver. In 1952, Bureau re- 
searchers reported a method for desorbing 
gold and silver from loaded carbon (3), 
Precious metals were stripped by contact- 
ing the carbon with a boiling NaOH-NaCN 
solution. With each pass of the solution 
through the carbon bed, some of the gold 
and silver was removed. After being 
passed through the carbon bed, the solu- 
tion entered an electrolysis cell, where 
the remaining gold and silver were depos- 
ited on the cathode. The barren solution 
was then reheated and recycled through 
the carbon. A 1 pet NaOH-0,1 pet NaCN 
solution at its boiling point desorbed 
more than 90 pet of the gold and silver 
in 4 to 6 h. The carbon could be reused 
up to 10 times without losing significant 
activity. 

Carbon adsorption-desorption-electro- 
winning permitted gold and silver 



recovery from slimy ores or the slimy 
portion of ores by eliminating the re- 
quirement of a clarified solution. This 
became the foundation for two important 
developments in gold and silver ore pro- 
cessing: carbon-in-pulp (CIP) cyanida- 
tion and heap leaching. The cylindrical 
electrolytic cell originally described 
(2^) is still used for electrowinning and 
is commonly referred to as a "Zadra" 
cell, even though many units in commer- 
cial use have been modified. Rectangular 
electrowinning cells are also used, to 
make better use of floor space than cyl- 
indrical cells. The NaOH-NaCN electro- 
lyte and steel wool cathode remain common 
to all gold-silver electrowinning cells. 
The Bureau's desorption-electrowinning 
process, known as the Zadra process, has 
been utilized in cyanide milling for more 
than 30 yr. 

Although gold and silver were eluted 
from the fruit pit carbon in 4 to 6 h 
in pilot-scale tests, this was not the 
general case. Commercial practice, which 
had adopted harder carbons made from 
coconut shells, showed that 24 to 48 h 
was required to desorb more than 90 pet 
of the precious metals using alka- 
line cyanide solution heated to boil- 
ing. Although this was better than not 
being able to strip the carbon, the long 
stripping time was undesirable. Bureau 
research was directed toward ways to 
decrease the stripping time. Two meth- 
ods, pressure stripping and alkaline- 
alcohol stripping, were developed; both 
greatly decreased the desorption time. 



PRESSURE STRIPPING AND ALKALINE-ALCOHOL STRIPPING 



In 1973, Bureau researchers showed that 
by using a pressure vessel and increasing 
the temperature of the stripping solu- 
tion, gold could be stripped from carbon 
in 2 to 6 h (4^) , The loaded carbon was 
conditioned with caustic-cyanide solution 
at 90° C and eluted with water at 150° C, 



One advantage was that consumption of 
cyanide and caustic was less with heated 
pressure stripping than it was using pro- 
longed ambient pressure stripping. The 
stripping solution was cooled to 90° C 
and the gold recovered by electrowinning. 



36 



In the alkaline-alcohol stripping meth- 
od, developed by Bureau researchers in 
1976, ambient pressure was used, but the 
modified stripping solution contained 20 
pet ethanol in addition to the alkaline 
cyanide (_5 ) . At 80° C, gold and silver 
were desorbed in 6 h. A concurrent de- 
velopment was the separation of gold from 
silver by precipitating silver as a sul- 
fide (6^). The separation of silver as a 
sulfide takes place after desorption from 



the carbon and results in a purer gold 
bullion. For pregnant solutions contain- 
ing considerable amounts of silver, the 
preferable sequence is to precipitate the 
silver before loading the carbon (_7 ) • 
A large carbon inventory is avoided by 
maintaining capacity for only gold ad- 
sorption. The Ag2S precipitate can be 
smelted to a silver bullion. The gold in 
the filtrate from silver precipitation is 
adsorbed, desorbed, and electrowon. 



SUMMARY 



Pressure stripping, alkaline-alcohol 
stripping, and ambient pressure stripping 
are all used by industry to strip gold 
and silver from activated carbon. Using 



any of these three methods, the carbon 
can be reused many times. The preference 
of the mill operator is the determining 
factor in choosing the stripping system. 



REFERENCES 



1. McQuiston, F. W. , Jr., and R. S. 
Shoemaker. Gold and Silver Cyanidation 
Plant Practice. AIME, v. 1, 187 pp., 
1975; V. 2, 263 pp., 1980. 

2. Zadra, J. B. A Process for the Re- 
covery of Gold From Activated Carbon by 
Leaching and Electrolysis. BuMines RI 
4672, 1950, 47 pp. 

3. Zadra, J. B., A. L. Engel, and 
H. J. Heinen. Process for Recovering 
Gold and Silver From Activated Carbon by 
Leaching and Electrolysis. BuMines RI 
4843, 1952, 32 pp. 

4. Ross, J. R. , H. B. Salisbury, and 
G. M. Potter. Pressure Stripping Gold 
From Activated Carbon. Pres. at Soc. 
Min. Eng. AIME Annu. Conf., Chicago, IL, 
Feb. 26-Mar. 1, 1973, 15 pp.; available 



upon request from Jean Beckstead, Bureau 
of Mines, Salt Lake City, UT. 

5. Heinen, H. J. , D. G. Peterson, and 
R. E. Lindstrom. Gold Desprption From 
Activated Carbon With Alkaline Alcohol 
Solutions. Ch. 33 in World Mining and 
Metals Technology, ed. by A. Weiss (Proc. 
Joint Min. and Metall. Inst, of Japan- 
AIME Meeting, Denver, CO, Sept. 1-3, 
1976). AIME, 1976, pp. 551-564. 

6. . Processing Gold Ores Using 

Heap Leach-Carbon Adsorption Methods. 
BuMines IC 8770, 1978, 21 pp. 

7. . Silver Extraction From Mar- 
ginal Resources. Pres. at 104th TMS-AIME 
Annu. Meeting, New York, Feb. 16-20, 
1975, 14 pp.; available upon request from 
J. A. Eisele, Bureau of Mines, Reno, NV. 



37 



HEAP LEACHING 
By J. A. Eiselel 



ABSTRACT 



The Bureau of Mines began investigat- 
ing heap leaching in 1969 in an effort 
to provide a low-cost means for recover- 
ing precious metals from ores that were 
too low in grade to be economically pro- 
cessed by conventional cyanide technol- 
ogy* By 1979, the Bureau had developed 



an agglomeration pretreatment method that 
permitted heap leaching to be applied to 
clayey and finely divided materials. To- 
day, heap leaching is widely used by the 
mining industry. This paper briefly de- 
scribes and discusses heap leaching and 
agglomeration pretreatment. 



INTRODUCTION 



Heap leaching was first used on oxide 
copper ores and uranium ores (JL^).^ It 
had the advantages of very low capital 
cost, low operating costs, and opera- 
tional flexibility. Bureau researchers 
proposed heap leaching in 1969 as a low- 
cost means for recovering gold values 
from disseminated gold ores with porous 
gangue, mine stripping waste, and submar- 
ginal ores (2-4). Heap leaching has 



since become an important factor in pre- 
cious metals recovery because it permits 
utilization of lean ores and wastes that 
cannot be economically processed by con- 
ventional agitation cyanidation. Fifteen 
years after the method's introduction, 
there are between 50 and 100 commercial 
gold and silver heap leaching operations, 
ranging in size from 10 st/wk to 10,000 
st/d. 



LEACHING 



In heap leaching, the crushed ore mate- 
rial is piled in heaps on impervious 
pads. A dilute alkaline-cyanide solution 
is distributed on top of the heap by a 
sprinkling system. The solution perco- 
lates through the heap and drains from 
the impervious pad. The pregnant gold 
solution from the heap typically contains 
from 1 to 3 ppm Au.- 

For many operations, precipitation of 
the metal values with zinc is not the 
most efficient recovery method. Another 
method, carbon adsorption, is very effi- 
cient for recovering of metals from di- 
lute solutions. The pregnant solution is 
passed through a series of columns con- 
taining activated carbon, and the gold is 
adsorbed. The resulting barren solution 

-I , 

'Supervisory chemical engineer, Reno 

Research Center, Bureau of Mines, Reno, 
NV. 

■^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



is fortified with reagents and recycled 
to the heap. Leaching continues until 
the gold extraction is completed. The 
gold-loaded carbon is stripped by carbon 
adsorption or zinc precipitation. Some 
pregnant solutions from silver heap 
leaching contain enough silver, typically 
10 to 20 ppm Ag, that zinc precipitation 
can be used to recover the silver. 

For heap leaching to be feasible, .the 
ore must be porous and permeable to the 
leaching solution. Since the ore is not 
finely ground, the cyanide ions must dif- 
fuse through the host rock to dissolve a 
gold particle. The dissolved gold must 
diffuse outward. This requires leaching 
cycles weeks or months long, and may re- 
sult in dissolved gold cyanide complexes 
that are not completely washed from the 
heap. Runoff from abandoned heaps is not 
discharged to surface or groundwater 
sources until the effluent from the spent 
heap is free of cyanide. A flow diagram 
of a typical heap leaching operation is 
shown in figure 1. 



38 



PRETREATMENT 



Although application of the heap leach- 
ing technique to gold and silver ores 
permitted development of many low-grade 
and/or small properties that otherwise 
would not have been exploitable, some 
ores were untreatable by heap leaching. 
This was due to two conditions: (1) The 
ore contained clay, which swelled on con- 
tact with the leaching solution, blocked 
the voids in the heap, and thereby pre- 
vented solution flow; and (2) the ore, 
after crushing to the liberation size for 
gold-silver extraction, generated an un- 
usually large amount of fines (minus 200 
mesh) that were washed into the voids by 



Makeup NaCN, 
CaO, HgO 



_.i__ 



Ore- 





Heap leaching 


--i- 


Car 


bon 












adsorption 


. 












' 








Carbon 
desorption 


Ttiermal 
reactivation 


r -*■ 




' 


keup No 


1 


i 

t 




1 




1 


Electrowinning 














Refining 





Gold-silver bullion 
FIGURE 1. - Flow diagram of heap leaching. 



the percolating leaching solution, caus- 
ing channeling and incomplete leaching 
of the gold and silver from the heap 
material. 

In 1979, Bureau researchers published a 
report describing an agglomeration method 
that was successful in overcoming these 
problems (_5-^) , Agglomeration as a pre- 
treatment for heap leaching consists of 
(1) mixing the crushed ore with portland 
cement, which acts as a binding agent, 
and lime to provide alkalinity; (2^) wet- 
ting the mixture evenly with solution, 
which may contain cyanide to start leach- 
ing before the heap is built; and (3) me- 
chanical tumbling of the mixture so 
the fine particles adhere to the larger 
particles. Several hours of aging are 
needed for the cement to bond the parti- 
cles. When stable bonds are formed, the 
agglomerates are very durable and resist- 
ant to degradation. This simple pre- 
treatment has increased the flow of 
some ores through the columns as much as 
6,000-fold and, in actual heaps, has 
decreased the leaching cycle to days in- 
stead of weeks. It is estimated that 
half of the heap leaching operations use 
some type of agglomeration pretreatment 
(7^), Agglomeration pretreatment can also 
be applied to finely divided material 
such as tailings or finely ground ore 
(8^) , The conditions required to form 
good agglomerates are more rigorous, but 
when the criteria are met, it is possible 
to process finely divided low-grade and/ 
or small resources. 



REFERENCES 



39 



1. Potter, G. M. Design Factors for 
Heap Leaching Operations, Min, Eng. , 
Mar. 1981, pp 277-281. 

2. Heinen, H. J., and B. Porter. Ex- 
perimental Leaching of Gold From Mine 
Waste. BuMines RI 7250, 1969, 5 pp. 

3. Merwin, R. W. , G. M. Potter, and 
H. J. Heinen. Heap Leaching of Gold Ores 
in Northwestern Nevada. (Pres. at AIME 
Annu. Meeting, Washington, DC, Feb. 16- 
20, 1969). Soc. Min. Eng. AIME preprint 
69-AS-79, 1969, 15 pp. 

4. Potter, G. M. Recovering Gold From 
Stripping Waste and Ore by Percolation 
Cyanide Leaching. BuMines TPR 20, 1969, 
5 pp. 

5. Heinen, H. J., G. E. McClel- 
land, and R. E. Lindstrom. Enhancing 



Percolation Rates in Heap Leaching of 
Gold-Silver Ores. BuMines RI 8388, 1979, 
20 pp. 

6. McClelland, G. E., and J. A. 
Eisele. Improvements in Heap Leaching To 
Recover Silver and Gold From Low-Grade 
Resources. BuMines RI 8612, 1982, 26 pp. 

7. McClelland, G. E., D. L. Pool, and 
J. A. Eisele. Agglomeration-Heap Leach- 
ing Operations in the Precious Metals In- 
dustry. BuMines IC 8945, 1983, 16 pp. 

8. McClelland, G. E., D. L. Pool, 
A. H. Hunt, and J. A. Eisele. Agglomera- 
tion and Heap Leaching of Finely Ground 
Precious-Metal-Bearing Tailings. BuMines 
IC 9034, 1985, 11 pp. 



40 



THE CARBON-IN-PULP PROCESS 
By Stephen D. Hill'' 



ABSTRACT 



This paper briefly reviews the develop- 
ment of the carbon-in-pulp (CIP) process 
for recovering gold and silver from cya- 
nide leach solutions. A brief history 
and a description of the various steps 
in the process are presented. Reference 



is made to published research findings, 
from both the Bureau of Mines and other 
sources, that contributed significantly 
to commercial development and utilization 
of the process. 



INTRODUCTION 



The application of activated carbon to 
recover gold and silver from cyanide 
leach solutions was patented as early as 
1894 (J_).2 T. G. Chapman, of the Univer- 
sity of Arizona, is credited with the 
original development of CIP processing in 
the late 1930' s. In Chapman's system, 
the dissolved gold was adsorbed by finely 
ground activated charcoal. Subsequently, 
the charcoal was separated from the pulp 
by flotation, and gold was recovered from 
the charcoal by burning or smelting (2). 

In an alternative system. Chapman used 
larger particles of activated carbon, 
particles that were much coarser than the 
ground ore, as the gold adsorbent. The 
carbon, enclosed in a cylindrical screen 
basket, was circulated and rotated within 
the leach pulp. A concentrated gold so- 
lution was obtained by stripping the 
gold-loaded carbon with hot cyanide solu- 
tion (_3 ) . Bureau of Mines researchers 
further modified the Chapman system in 
the early 1950' s by adding an electro- 
lytic cell for continuous removal of gold 
from the hot strip liquor (4). 

A small commercial CIP plant was op- 
erated from 1954 to 1960 at the Golden 
Cycle Corp.'s Carlton Mill in Cripple 
Creek, CO. In that operation, coarse 
carbon in screen baskets was loaded 

'Research Director, Salt Lake City Re- 
search Center, Salt Lake City, UT. 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



with gold 
stripped 
method (5) 
In 1970, 
continued 



to about 40 tr oz/st Au, then 
and recycled by the Zadra 



the Homestake Mining Co. dis- 
mercury amalgamation in its 
milling plant in Lead, SD, because of 
downstream pollution. Since the ore had 
a high slime content, Homestake' s diffi- 
culties provided an opportunity for the 
Bureau to test its own concepts for im- 
proved gold extraction, loading, and car- 
bon handling in a modified cyanide CIP 
system. Research in these areas was per- 
formed under a cooperative agreement be- 
tween the Bureau and Homestake. At the 
time, conventional plants used a costly 
system for processing low-grade, lode 
gold ores that contained finely dissemi- 
nated gold. Considerable capital invest- 
ment was required for fine grinding, 
agitation leaching in cyanide solution, 
countercurrent decantation in thickeners 
for solid-liquid separation, residue 
washing, clarification of pregnant solu- 
tion by filtering, vacuum deaeration, and 
precipitation of the gold by reacting the 
solution with zinc powder. 

The alternative system, proposed by 
the Bureau, optimized the following oper- 
ations: pulp leaching, CIP loading, 
carbon separation, stripping, electrowin- 
ning, and carbon regeneration. The inno- 
vative system was perfected and demon- 
strated at Homestake in a miniplant that 
handled 2 st/d of slime pulp. By extrap- 
olation of the design from the mini- 
plant, Homestake was able to start its 



41 



commercial CIP plant, treating 2,000 st/d 
slime, in 1973 (6^). Bureau metallurgists 
reported on the Homestake project and the 
new CIP process at the American Mining 
Congress Mining Convention in Denver, CO, 
the same year (7^) . 

The Bureau subsequently developed the 
carbon pressure stripping process, and 
refinements were made in a new design for 
the electrowinning cell. Cost data pre- 
sented in 1974 clearly showed that the 
CIP system was superior to the conven- 
tional countercurrent decantation system 
(8^). The MINTEK organization in Johan- 
nesburg, Republic of South Africa, began 
work on CIP in 1976, and many significant 
developments have resulted from its in- 
tensive efforts (9) . 

Other CIP plants patterned after the 
Bureau process include Duval Corp.'s Bat- 
tle Mountain, NV, operation in 1978 (10); 
Pinson Mining Co.'s Winnemucca, NV, plant 
in 1980 (11); and Freeport Gold Co.'s 
operation near Elko, NV, in 1981, the 
largest plant of its kind in the United 
States (12) . Getty Mining Co., Mercur, 
UT, started a carbon-in-leach (CIL) 
plant, an offshoot from the CIP process, 
in 1983 (13). Today, the CIP process 
is the standard of the gold industry 
throughout the world, with 30 plants in 
North America, 10 plants in Australia, 
and approximately 20 plants in the Repub- 
lic of South Africa using the process 
(24-J^) . 

LEACHING 

The CIP process, shown in figure 1, 
Includes cyanide agitation leaching, 
countercurrent carbon-pulp contact, and 
carbon-pulp separation. The leaching 
step of the process, usually accomplished 
in a multistage system, comprises sev- 
eral tanks; four to six stages are typi- 
cal. The ore, normally ground to minus 
100 mesh or finer, is leached in a thick 
slurry of about 45 to 50 pet solids. 
Lime or caustic is added to maintain 
protective alkalinity. To speed up gold 
dissolution, the pulp is vigorously aer- 
ated and agitated during leaching. The 
retention time necessary for complete 



Gold ore 

Crushing 

Grinding -• — NaCN plus lime 

45-50 pet solids pulp, minus 100 mesh or finer 




Leach agitators 




Stripping »• Electrowinning ►Gold 



Carbon reactivation 



Makeup 

minus 10- plus 20-mesh 

activated c arbo n 

1 



FIGURE 1.- CIP process. 

dissolution of the gold depends greatly 
on the particle size of the gold, but 2 
to 24 h is typical. Other factors that 
influence the rate of dissolution include 
particle size of the ore, clay content, 
pulp density, pH, cyanide strength, tem- 
perature, and slurry viscosity. 

The Dorr agitator, with a slow-speed 
center sweep and either peripheral or 
center-column airlifts, has worked well 
on finely ground ores. Pachuca tanks 
also have been successful; they may be 
capable of handling a coarser feed than 
the traditional Dorr tanks, A draft 
tube-type agitation tank has also been 
used successfully. The type of vessel 
most suitable for CIP processing depends 
on the type of ore being treated and the 
prevailing operating conditions. In most 
cases, a deep tank with turbine-type me- 
chanical agitators and low tip speeds is 
preferred (17), Dissolution of gold and 
silver is essentially completed before 
the leached pulp moves to the CIP adsorp- 
tion circuit. 



42 



ADSORPTION 



CARBON SEPARATION 



The CIP adsorption circuit is a cascade 
of agitated vessels through which the 
pulp flows by gravity. In each vessel, 
the pulp is contacted with carbon gran- 
ules that preferentially adsorb gold and 
silver from the solution as the pulp 
overflows from one vessel into the next. 
Periodically, a portion of the carbon in- 
ventory is transferred up the cascade to 
the next vessel by airlifting a quantity 
of carbon-bearing pulp. The transfer 
system is abrasive; therefore, coconut 
shell activated carbon is generally used 
because it is resistant to abrasion. 
There is some variation in the size of 
carbon particle used. The only strict 
criterion for carbon size is that parti- 
cles be large enough to allow easy sepa- 
ration from the pulp; a minus 6- plus 16- 
mesh carbon size is quite common. 

The optimum loading of gold and silver 
on the carbon is generally a matter of 
economics and is very dependent on metal 
prices. Consideration must be given to 
the ore grade, solution value, gold in- 
ventory, and security. Frequent strip- 
ping and handling can result in lower 
loading, so most large operations usually 
load the carbon to 200 to 400 tr oz/st Au 
before the carbon is removed from the 
first tank in the cascade. 

Several factors influence the adsorp- 
tion capacity of gold onto activated car- 
bon from cyanide solutions. Some of the 
more important are the ionic strength and 
pH value of the solution, temperature, 
the presence of competing metal ions and 
poisons, and the nature of the carbon. 
In addition, a number of factors influ- 
ence the extraction rate of gold cyanide 
from activated carbon in the stirred 
tanks, the most important of which are 
the mixing efficiency in the tanks, pulp 
viscosity, and carbon granule sizes (18) , 

The Bureau of Mines recommends that ad- 
sorption efficiencies be determined by 
conducting laboratory CIP batch tests to 
obtain adsorption rate data and establish 
equilibriim isotherms (19-20) . These ad- 
sorption rates and equilibrium curves can 
then be used to design the optimum ad- 
sorption system. 



Methods of separating the loaded carbon 
particles from the gold-depleted pulp are 
numerous and widely varied. In early CIP 
circuits, screen baskets or external vi- 
brating screens were used extensively. A 
vibrating screen with a minus 20- plus 
24-mesh stainless steel square mesh deck 
has proven to be reliable. Such screens 
have worked well at a number of plants 
without malfunctioning. 

A problem of carbon fines being gen- 
erated caused operators to look for im- 
proved methods that would save power, re- 
duce fine carbon generation, and reduce 
capital costs. The new concepts that re- 
sulted include either a peripheral screen 
or a submerged launder-type screen that 
fits across the top of the CIP tank. 
Typically there are two or three such 
launders on each tank. Each launder has 
two or three removable screens set in the 
side panels, which may be vibrated for 
carbon transfer, if necessary, 

CARBON-IN-LEACH 

Leaching generally requires a much 
longer pulp residence time than adsorp- 
tion. Consequently, it is possible to 
reduce the equipment requirement in the 
CIP circuit by using the leach vessels 
for both cyanidation and adsorption si- 
multaneously, thus eliminating the need 
for a separate adsorption cascade. Such 
a system is called a carbon-in-leach 
(CIL) circuit (21) . Because the acti- 
vated carbon adsorbs gold and silver from 
solution, not from the ore, there are ad- 
vantages in partially preleaching the 
ore before adsorption starts. Leaching 
can go on in the presence of carbon that 
is moving countercurrent to the flow of 
pulp, similar to the operation of a stan- 
dard CIP circuit, 

RECOVERY OF GOLD AND SILVER 

The activated carbon, loaded with 
gold and/or silver, may be shipped to a 
smelter to recover the precious metal 
values. Such a procedure is sometimes 
used, particularly for small mines whose 



capital is limited and reserves are 
small. However, for better economy in 
larger operations, the carbon is prefer- 
ably stripped, reactivated, and re- 
used in the process. Details of carbon 



43 



stripping, regeneration, and recovery of 
precious metals by electrowinning from 
strip solution are discussed in other pa- 
pers in this Information Circular. 



REFERENCES 



1. Johnson, W. D. Abstraction of 
Gold and Silver From Their Solutions in 
Potassium Cyanide. U.S. Pat. 533,260, 
May 18, 1894. 

2. Chapman, T. G. Cyanidation of 
Gold Bearing Ores. U.S. Pat. 2,147,009, 
Sept. 22, 1939. 

3. Crabtree, E. H. , Jr., V. W. Win- 
ters, and T. G. Chapman. Developments 
in the Application of Activated Carbon 
to Cyanidation. Metall. Trans., v. 187, 
Feb. 1950, pp. 217-222. 

4. Zadra, J. B., A. L. Engel, and 
H. J. Heinen. Process for Recovering 
Gold and Silver From Activated Carbon 
by Leaching and Electrolysis. BuMines 
RI 4843, 1952, 32 pp. 

5. Seeton, D. A. A Review of Carbon 
Cyanidization. Min. Mag., v. 51, July 
1961, pp. 13-15. 

6. Hall, K. B. Homestake Uses Car- 
bon-in-Pulp To Recover Gold From Slimes. 
World Min., v. 27, No. 12, 1974, p. 44. 

7. Potter, G. M. , and H. B. Salis- 
bury. Innovations in Gold Metallurgy. 
(Pres, at Am. Min. Congr. Min. Conv. and 
Environ. Show, Denver, CO, Sept. 9-12, 
1973.) BuMines preprint (Salt Lake City, 
UT), 1973, 12 pp. 

8. Rosenbaum, J. B. Minerals Extrac- 
tion and Processing: New Developments. 
Science, v. 191, 1976, pp. 720-723. 

9. Laxen, P. A., G. S. M, Becker, and 
R. Rubin. Developments in the Applica- 
tion of Carbon-in-Pulp to the Recovery of 
Gold From South African Ores. J. S. Afr. 
Inst. Min. & Metall., v. 79, June 1979, 
pp. 315-326. 

10. Jackson, D. (ed.). How Duval 
Transformed Its Battle Mountain Proper- 
ties From Copper to Gold Production. 
Eng. and Min. J., v. 183, No. 10, 1982, 
pp. 95-99. 

11. Mining Magazine. Pinson, Nevada — 
New Open Pit Gold Mine. V. 145, July 
1981, p. 5. 



12. Jackson, D. (ed.). Jerritt Canyon 
Project. Eng. and Min. J., v. 183, July 

1982, pp. 54-58. 

13. Burger, J. (ed.). Mercur is Get- 
ty's First Gold Mine. Eng, and Min. J., 
V. 184, No. 10, 1983, pp. 48-51. 

14. Schreiber, H. W, , and M. E. Emer- 
son. North American Hardrock Gold Depos- 
its. Eng. and Min. J., v. 185, No. 10, 
1984, pp. 50-57. 

15. Todd, J. C. New Mines Fuel Aus- 
tralian Gold Boom, Eng. and Min. J. , v. 
185, No. 5, 1985, pp. 11-13. 

16. Danne, R. (MINTEK, Johannesburg, 
Republic of South Africa) . Private com- 
munication, 1985; available from S. D. 
Hill, Bureau of Mines, Salt Lake City, 
UT. 

17. Potter, G. M. , R. S. Shoemaker, 
K. B. Hall, and D. M. Duncan. Carbon-in- 
Pulp Processing of Gold and Silver Ores. 
Min. Eng. (Littleton, CO), pt. 1, v. 33, 
No. 9, 1981, pp. 1331-1335; pt. 2, v. 33, 
No. 10, 1981, pp. 1441-1444. 

18. Fleming, C. A. Recent Develop- 
ments in Carbon-in-Pulp Technology. Pa- 
per in Hydrometallurgical Research, De- 
velopment, and Plant Practices, ed. by 
K. Osseo-Asare and J. D. Miller. Metall. 
Soc. AIME, 1983, pp. 839-857. 

19. Hussey, S. J., H. B. Salisbury, 
and G. M. Potter. Carbon-in-Pulp Silver 
Adsorption From Cyanide Leach Slurries 
of a Silver Ore. BuMines RI 8268, 1978, 
22 pp. 

20. . Carbon-in-Pulp Gold Ad- 
sorption From Cyanide Leach Slurries. 
BuMines RI 8368, 1979, 22 pp. 

21. Newrick, G. M. , G. Woodhouse, and 
D. M. G. Dods. Carbon-in-Pulp Versus 
Carbon-in-Leach. World Min., v. 36, June 

1983, pp. 48-51. 



44 



PRECIOUS METALS RECOVERY FROM ELECTRONIC SCRAP AND SOLDER 
USED IN ELECTRONICS MANUFACTURE 



By B. W, Dunning, Jr. "I 



ABSTRACT 



Electronic scrap from obsolete and/or 
damaged avionics or from manufacturing 
sources poses a problem for the owner 
or generator of the scrap in terms of 
its fair value. The complexity of this 
material, as well as the low concentra- 
tion of precious metals (generally less 
than 1 pet) , makes it difficult to obtain 
a representative sample for assay. The 
owner or generator is therefore dependent 
on the reliability and competence of a 
toll refiner to obtain a fair value. 



The Bureau of Mines has investigated var- 
ious procedures for either concentrating 
precious metals from electronic scrap in- 
to an easily assayable form or for re- 
covering fairly pure gold, silver, or 
platinum-group metals (PGM). Procedures 
the Bureau has studied are described 
and discussed in this paper, including 
hand dismantling, mechanical processing, 
pyrometallurgy , hydrometallurgy , and 
electrometallurgy. 



INTRODUCTION 



Gold, silver, and PGM are widely used 
in electronic and electrical components 
to provide long-term reliability. Fabri- 
cation of equipment used by the military 
consumes the largest portion of precious 
metals used in the electronics and elec- 
trical industry. The disposal of obso- 
lete and/or damaged military electronics 
(reportedly totaling more than 10,000 
st/yr) at a fair value is a pressing 
problem for the Department of Defense 
(DOD) . This is partly due to the highly 
variable and complex nature of military 
electronic scrap, which makes it diffi- 
cult to obtain a homogeneous sample for 
precious metals analysis. Private indus- 
try has the same problems, but faces few- 
er constraints in dealing with them. 

The Bureau of Mines, in cooperation 
with DOD, the National Association of Re- 
cycling Industries (NARI) , and others. 



has developed many procedures for deter- 
mining precious metals content in scrap. 
However, some of these procedures are not 
economical except under special circum- 
stances. Initial research to determine 
precious metals content in electronic 
scrap relied on hand dismantling and 
object recognition. Subsequent studies 
investigated mechanical processing as a 
method for obtaining metal concentrates 
containing the major portion of the pre- 
cious metals. In later studies, these 
precious metal concentrates were treat- 
ed using various procedures, includ- 
ing pyrometallurgy, hydrometallurgy, and 
electrometallurgy, to further concentrate 
or recover the precious metals. These 
procedures are discussed in the following 
sections, and data concerning the effec- 
tiveness of each procedure are presented. 



HAND CHARACTERIZATION OF SELECTED AVIONIC MODULES (1-2)2 



In 1977, the Defense Property Disposal 
Service (of the DOD) initiated a test re- 
covery program that included having the 

'Supervisory metallurgist, Avondale Re- 
search Center, Bureau of Mines, Avondale, 
MD. 



Bureau of Mines hand-process a controlled 
sample lot containing a variety of "black 
boxes" (surplus electronic units so 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



45 



designated because of their black out- 
er cases). The program was to include 
determination of base and precious met- 
als content, removal of salable modules, 
and a monetary evaluation of each box. 
Several identical lots were prepared 
with components intact to assure a basis 
for direct comparison. The Bureau re- 
ceived one lot of boxes for hand dis- 
mantling to determine the potential yield 
of base and precious metals from each 
unit. An identical lot of boxes was 
placed on display at the Defense Con- 
struction Supply Center, Columbus, OH, 
for private industry to bid on determina- 
tion of the following: 

1. Relative cost-effectiveness of and 
estimated cost per item for currently 
practicable methods for precious metals 
segregation, identification, and recovery 
and for serviceability testing of usable 
modules from individual boxes, 

2. Reutilization and sales potential 
of usable modules and components, with 



sales potential supported by quotations 
from prospective purchasers. 

All interested industry groups indi- 
cated that the sample lot was too low 
in value to warrant bidding and too 
old for reutilization of usable mod- 
ules. More than half of the sample con- 
sisted of radio receivers, transmitters, 
tuners, and power supplies; the remain- 
der consisted of miscellaneous naviga- 
tional and communication equipment. All 
units appeared to have been produced 
prior to 1957 and did not contain any 
printed circuits. The weights for in- 
dividual units in the sample lot ranged 
from 2-1/2 to 58 lb. The total weight 
of the sample lot was 726 lb. The avi- 
onlc units came from the military air- 
craft storage located in the Arizona 
desert near Tucson. A summary of the 
materials composition of the 36 hand- 
dismantled avionic units is listed in 
table 1. 



MECHANICAL PROCESSING OF ELECTRONIC SCRAP AND GOLD AND SILVER 
DISTRIBUTION IN THE VARIOUS FRACTIONS (3-5) 



Approximately 5 st of general avionic 
scrap "black boxes" was mechanically pro- 
cessed through a series of unit opera- 
tions. These unit operations included a 
hammer mill, air classif ier-baghouse, 
magnetic separator-trommel, vibrating 
screen, rolls crusher, wire separator 
screen, magnetic precleaner, eddy-current 
separator, and high-tension separator. 
The various materials recovered from 
these unit operations and their distribu- 
tion are listed in table 2, along with 
the assay and the amount of gold and sil- 
ver in each material fraction. Three ma- 
terial fractions, the lights, wire bun- 
dles, and the metallics from high-tension 
separation, contained most of the gold 
and silver. It was determined that these 
materials, representing 34 pet of the 
original 5 st, could be economically pro- 
cessed by a toll refiner to obtain credit 
for the precious metals. 



Hand-segregated printed cards and elec- 
trical plugs and connectors were also 
processed through a modified series of 
mechanical unit operations. Rolls crush- 
ing and magnetic precleaning were not 
deemed necessary for processing these 
items. The amount and distribution of 
materials recovered from circuit cards 
using the various unit operations are 
shown in table 3, along with the assay 
and the amount of gold and silver in each 
material fraction; the results shown are 
from processing approximately 920 lb of 
printed-circuit cards. 

The amount and distribution of materi- 
als recovered from mechanically process- 
ing some 600 lb of electrical plugs and 
connectors are listed in table 4, along 
with the assay and the amount of gold and 
silver in these components. Most of the 
gold and silver was found in the high- 
tension metal and baghouse lights. 



46 



TABLE 1. - Summary of materials composition of hand-dismantled avionics units, 
weight percent 



Uniti 



Aluminum 
base 



Copper 
base 



Magnetic 
metals 



Stainless 
steel 



Nonmetals 



Fraction con- 
taining pre- 
cious metals^ 



Receiver-transmitter 

Tuner, radio 

Tuner, radio 

Tuner, radio 

Radio receiver 

Converter 

Keyer 

Amplifier 

Video amplifier 

Video decoder 

Radio receiver 

Receiver-transmitter 
Control transmitter. 

Video coder 

Indicator 

Tuner, radio 

Tuner, radio 

Tuner, radio 

Tuner, radio 

Receiver 

Coder transmitter 

set 

Receiver-transmitter 
Azimuth indicator, . . 

Receiver 

Receiver-transmitter 

Indicator 

Azimuth indicator. . . 

Power supply 

Receiver-transmitter 

Power supply 

Power supply 

Power supply 

Storage unit 

N^ compass 

Inverter 

Electron tube 



20. 
57. 
56. 
54. 
25. 
47. 
28. 
32. 
29. 
51. 
28. 
56. 
32. 
54. 
39. 
59. 
62. 
61. 
51. 
34. 

45. 
36. 
36. 
25. 
32. 
43. 
40. 
27. 
28. 
6. 
28. 
25. 
58. 
16. 
32. 
14. 



35.8 
20.8 
20.1 
17.6 
28.3 
23.5 
23.6 
22.2 
23.5 
13.3 
25.9 
17.0 
22.4 
16.0 
12.8 
16.5 
12.4 
14.7 
17.7 
17.4 

15.3 
18.8 
19.8 
21.7 
17.7 
17.0 
17.5 
19.0 
18.1 
24.4 
16.8 
15.9 
14.1 

5.1 
19.4 

8.4 



19.1 

9.5 

9.1 

10.5 

29.0 

11.0 

21.0 

23.8 

19.3 

11.5 

23.5 

13.5 

12.9 

9.0 

12.9 

8.3 

8.3 

9.3 

14.1 

26.7 

22.5 
23.1 
21.1 
30.1 
21.2 
19.6 
20.8 
31.1 
28.6 
47.6 
34.5 
33.3 
15.7 
43.8 
41.4 
69.4 



4.3 
4.4 
5.3 
7.5 
2.3 
3.2 
5.7 
2.3 
8.9 
8.5 
1.5 
1.3 
13.7 
3.4 
3.9 
5.5 
8.5 
5.9 
7.9 
1.7 

1.4 
3.3 
4.4 
3.4 
3.4 
3.0 
4.3 
2.3 
4.0 
2.5 

.6 
3.7 

.5 
33.3 

.8 
5.5 



20.3 

7.9 

8.6 

9.5 

15.3 

14.8 

21.2 

18.7 

18.9 

15.1 

20.4 

11.5 

18.9 

17.3 

30.9 

10.1 

8.8 

8.3 

9.1 

19.6 

15.5 
18.3 
18.5 
18.9 
25.7 
16.5 
17.1 
19.7 
20.5 
19.2 
19.4 
21.9 
11.3 
1.7 
5.8 
2.3 



14.5 

13.5 

13.4 

11.6 

11.1 

8.2 

5.2 

5.2 

4.7 

3.6 

3.5 

2.8 

2.8 

2.8 

2.6 

2.5 

2.3 

2.3 

2.2 

2.2 

2.2 

1.9 

1.8 

1.8 

1.5 

1.4 

1.3 

1.3 

1.2 

1.1 

1.0 

.8 

.3 

.2 

.2 





'where units of the same kinc 
individual unit. 

^Percentage of the total 
hand segregation of silver- 



are listed more than once, each listing represents an 



black box we 
and gold-co 



ight that contained precious metals, based on 
ated components through visual examination. 



47 



TABLE 2. - Material distribution, concentration, and weight of contained precious 
metals from fractions of mechanically beneficiated general avionic scrap 



Fraction 


Weight, 
lb 


Distribution, 
wt pet 


Concentration, 
wt pet 


Contained precious 
metals, tr oz 




Au 


Ag 


Au 


Ag 


Baghouse lights 

Wire bundles ............. 


930 
860 

1,204 
1,825 

425 

1,161 

965 

552 
444 
773 

1,318 

23 


8.9 
8.2 

11.5 
17.4 

4.0 

11.1 

9.2 

5.3 

4.2 
7.4 

12.6 

.2 


0.021 
.021 

.015 
0) 

.064 

ND 

.003 

.099 
.074 
.031 

.017 


0.33 
.77 

.07 
(M 

.41 

.017 

.64 

2.01 
1.96 
1.09 

.30 


2.8 
2.6 

2.6 
0) 

4.0 
ND 
.42 

8.0 
4.8 
3.5 

3.3 


44.7 
96.6 


Magnetics: 

Minus 1/2 in 


12.3 


Minus 1 plus 1/2 in.... 
Eddy-current: 2 

Precleaner magnetics... 
Aluminum. .••.....•..••. 


0) 

25.4 
2.9 


Middles 


90.1 


High-tension metal: ^ 

Minus 1/4 in. .......... 


161.7 


Minus 1/2 plus 1/4 in.. 
Minus 1 plus 1/2 in.... 
High-tension rejects 

(minus 1 in) ^ 

Hammer-mill knockout box 
contents 


127.0 
122.9 

57.7 

(M 


Composite, total or 
average 


10,480 


100.0 


.021 


.48 


32.02 


741.3 



ND Not detected. 

'visual observation of this fraction showed no evidence of precious metals except 
for a few large power transistors. 

^Fractions obtained using ramp-type eddy-current separator. 

^Fractions obtained using high-tension separator. 

^Not determined. 



TABLE 3. - Material distribution, concentration and weight of contained precious 
metals in fractions of mechanically beneficiated printed-circuit cards 



Fraction 


Weight, 
lb 


Distribution, 
wt pet 


Concentration, 
wt pet 


Contained precious 
metals, tr oz 




Au 


Ag 


Au 


Ag 


Baghouse lights . . . . < 

Wire bundles 


214.6 

11.2 

283.6 

16.6 

324.9 

72.0 


23.2 

1.2 

30.7 

1.8 

35.2 

7.8 


0.075 
.068 
.030 

ND 

.150 

.023 


0.61 
.97 
.05 

.01 

1.55 

.37 


2.35 

.11 

1.24 

ND 

7.11 

.24 


19.1 
1.6 


Magnetics (minus 1/2 in). 
Eddy-current aluminum 

(minus 1/2 in) 

High-tension metal (minus 

1/2 in) 


2.1 
.02 
73.4 


High-tension rejects 
(minus 1/2 in) 


3.9 


Composite, total or 
average 


922.9 


99.9 


.082 


.74 


11.05 


100.1 



ND Not detected. 



48 



TABLE 4. - Material distribution, concentration, and weight of contained precious 
metals in fractions of mechanically beneficiated electrical plugs and connectors 



Fraction 


Weight, 
lb 


Distribution, 
wt pet 


Concentration, 
wt pet 


Contained precious 
metals, tr oz 




Au 


Ag 


Au 


Ag 


Baghouse lights 

Wire bundles 


81.6 

6.6 

53.2 

66.6 

199.8 

191.8 


13.6 
1.1 
8.9 

11.1 

33.3 

32.0 


0.081 
.043 
.020 

ND 

.137 

ND 


0.88 

1.34 

.10 

.015 
1.12 
ND 


0.96 
.04 
.16 

ND 

3.99 

ND 


10.5 
1.3 


Magnetics (minus 1/2 in). 
Eddy-current aluminum 

(minus 1/2 in) 

High-tension metal (minus 

1/2 in) 

High-tension rejects 

(minus 1/2 in)' 


.78 

.15 
32.6 

ND 


Composite 


599.6 


100.0 


.059 


.52 


5.15 


45.3 



ND Not detected. 

'visual observation of the high-tension rejects (minus 1/2 in) indicated that the 
shattered pieces of plastic, ceramic, hard rubber, and other types of insulator mate- 
rial contained no precious-metals-bearing contact pins after shredding. 



In processing either whole avionic 
units or hand-segregated circuit cards , 
electrical plugs, and connectors, the 
greatest amount of gold and silver is 
always concentrated in the high-tension 
metal product. Most of the remaining 
gold and silver concentrates in the wire 
and baghouse lights. These fractions are 
best handled by a toll refiner. A 
toll refiner will first incinerate these 



fractions separately to remove the or- 
ganics. After incineration, the high- 
tension metal product and wire are dis- 
solved in a molten heel of copper. The 
melt is then thoroughly mixed and cast 
into ingots for assay of the precious 
metals. The baghouse lights are also in- 
cinerated, then ball or rod milled; the 
resulting powder is then blended, coned, 
quartered, and assayed. 



HYDROMETALLURGICAL RECOVERY FROM ELECTRONIC SCRAP 



GOLD AND SILVER FROM HIGH-TENSION 
METALLIC FRACTION (6-7) 



three stages with 20-vol-pct H2 SO4 to re- 
move the copper. 



The high-tension metallics from me- 
chanically processed avionic units, rep- 
resenting about 17 pet of the units' 
total weight, contained slightly more 
than 50 pet of the total gold and sil- 
ver. The high-tension fraction, with 
its high surface-area-to-weight ratio, is 
ideal for upgrading by leaching. The 
elemental composition of this fraction is 
listed in table 5. The major base metals 
in the high-tension fraction were copper, 
aluminum, and iron, in that order. Alu- 
minum is removed by leaching with 20- 
wt-pct NaOH. After washing, the residue 
is countercurrently pressure leached in 



TABLE 5. - Elemental composition of 
high-tension metallic fraction 
from avionic scrap 



Ag. 
Al. 
Au. 
Cr, 
Cu. 



Cone, 

wt pet 



1.37 
27.2 
.12 
.1 
38.4 



Fe. 
Ni. 
Pb. 
Sn. 



Cone, 
wt pet 

9.4 

3.2 

.2 

2.1 



NOTE. — Remainder was mostly Si as Si02 
(fiberglass and plastic filler). 



49 



The distribution of metal values in the 
leaeh solutions and in the residues is 
listed in table 6. Copper is reeovered 
from the pregnant leaeh liquor by eemen- 
tation with transformer steel from the 
coarse magnetic fraction of mechanically 
processed avionics. The cement copper 
is a good grade and assays better than 
90 pet. 

TABLE 6. - Concentration and distribu- 
tion of metal values in leaeh solu- 
tions and residues 





Leach solutions 


Residues 


Elements 


Cone, 


Distri- 


Cone, 


Distri- 




g/L 


bution, 
pet 


wt pet 


bution, 
pet 


Ag 


<0.001 


<0.004 


8.13 


99.1 


Au 


ND 


ND 


.74 


>99.9 


Cr 


.0004 


6.2 


.42 


94.0 


Cu 


36.3 


89.3 


24.4 


10.9 


Fe 


2.9 


29.0 


39.6 


71.0 


Ni 


1.02 


38.8 


9.0 


61.1 


Pb 


.003 


1.4 


1.1 


98.6 


Pd 


ND 


ND 


.21 


99.9 


Sn 


.14 


7.9 


9.5 


91.1 



ND Not detected. 

Silver is reeovered from the pressure- 
leached residue by dissolution in 50- 
vol-pet HNO3 and subsequent precipitation 
with NaCl as AgCl. This AgCl is reduced 
to silver metal by mixing with Na2C03 and 
heating to 600° C. Copper not extracted 
during H2SO4 leaching is reeovered from 
the silver-free HNO3 leaeh solution by 
cementation with steel scrap. Gold is 
extracted from the HNO3 leaching residue 
with aqua regia and precipitated with 
NaHS03. Analyses of the gold and silver 
products and the final residue is listed 
in table 7. The final residue consisted 
mostly of acid-insoluble noncombustibles 
(mostly silica) , stainless steel, and 
tin. 

PLATINUM-GROUP METALS (_8) 

Hydrometallurgical techniques have al- 
so been investigated as a means for 
recovering platinum-group metals from 
electronic scrap. 



Telephone Relay Scrap 

NARI estimates that about 1,000 st/yr 
of telephone relay scrap is available for 
processing. Palladium, gold, and plati- 
num are found in the contact points, 
which are brazed to cupronickel wires. 
These contact points represent approxi- 
mately 0.06 pet of the modular weight. 
Mechanical processing, including shred- 
ding, air classification, magnetic sep- 
aration, screening, and high-tension sep- 
aration (HTS), produced an HTS metallic 
fraction containing most of the relay 
contact points. A portion of this HTS 
concentrate representing 8.7 pet of the 
original scrap was pressure leached in 
stainless steel autoclaves in a two-stage 
countercurrent arrangement with solid- 
liquid separation between each stage. 
The purpose of the first-stage leaeh was 
to react fresh HTS concentrate with the 
second-stage liquor to produce a pregnant 
liquor with a low concentration of H2 SO4 
(less than 10 g/L). Partially leached 
solids from the first-stage leach were 
fed to the second stage, where they were 
leached with a fresh H2SO4-HNO3 solution. 
Measured quantities of HTS-PGM con- 
centrate (usually 120 g), 1 L of H2 SO4 
solution (20 vol pet), and HNO3 (20 mL 
to 120 g concentrate) were charged to 
the autoclave and heated to 90° C with a 
lOO-lbf/in^ air overpressure. Air was 
sparged through the autoclave (about 10 
bubbles per second) during all leaches. 
After leaching, the slurry was removed 
from the autoclave and filtered, and the 
products were analyzed. Increasing the 
reaction time from 1 to 4 h increased 
base metal extraction from 96 to 99 pet. 
Quantitative analyses of the residue and 
filtrate are listed in table 8. The res- 
idue, representing 0.2 pet of the HTS-PGM 
concentrate, was suitable for shipping to 
a toll refiner for recovery of silver, 
gold, and PGM. 

Reed Switches 

NARI sources estimate that approximate- 
ly 10 St of reed switches, a valuable 
scrap material, is available for process- 
ing annually. Reed switches consist of 



50 



TABLE 7. - Semiquantitative spectrochemical analyses of products 
from pressure-leached residue, weight percent 



Element 


Pro 


duct 


Final residue 




Gold 


Cement silver^ 




Ag 

Al 


0.03 - 0.3 
.003- .03 

>10 
.003- .03 
.003- .03 
.03 - .3 
.003- .03 
.01 - .1 

1 - 10 
.1 - 1 
.03 - .3 


>10 
0.03 - .3 
.0003- .003 

ND 
.03 - .3 
.03 - .3 
.003 - .3 
.3-3 

ND 
.1 - 1 
.3-3 


0.0003- 0.003 
.3-3 


Au 


ND 


Cr 


.3-3 


Cu 


.1-1 


Fe 


.3-3 


Ni 


.01 - .1 


Pb 


.1-1 


Pd 


ND 


Si 


>10 


Sn 


1 - 10 



ND Not detected. 

^Quantitative analysis of remaining cement silver after semi- 
quantitative spectrochemical analysis, in weight percent: 97.5 
Ag, 0.028 Al, 0.53 Cu, 0.12 Pb, and 1.82 Sn. 



TABLE 8. - Quantitative analyses of 
residue and filtrate from leaching 
telephone relays 



Element 


Concentration 




Residue, wt 


pet 


Filtrate, g/L 


Ag 


1.3 




0.001 


Au 


14.8 




.0004 


Cu 


1.3 




67.9 


Cr 


<.003 




.0003 


Mn 


ND 




.03 


Ni 


.09 




14.4 


Pd 


77.7 




.0007 


Pt 


.29 




.009 



ND Not detected. 

two magnetizable reeds with their extrem- 
ities fused in a glass envelope. The in- 
ner ends of the reeds are plated with 
gold and then rhodium. A simple rolls 
crushing step followed by magnetic sepa- 
ration of the reeds from the glass powder 
was the only preprocessing necessary to 
prepare the reed switches for hydrometal- 
lurgical treatment. Assay of the reed 
switches showed them to be, in weight 
percent, 49 Co, 48 Fe, 2 V, 0.5 Au, and 
0.4 Rh. 

The approach to recovering gold and 
rhodium from this scrap was to dis- 
solve the cobalt-iron substrate in either 
HCl (1:1) or HNO3 (1:2) using ultrasonic 
agitation. Dissolution of the base-metal 



substrate was relatively fast except for 
the working face of the reed plated with 
gold and rhodium. Complete removal of 
base metals required 12 h of continuous 
leaching. The insoluble residue was a 
fine gold-rhodium sand assaying approxi- 
mately 50 pet Au and 50 pet Rh. A spec- 
trochemical analysis of the gold-rhodium 
sand is listed in table 9. Gold was 
readily removed from the Rh with aqua 
regia, leaving a pure 99.8-pct-Rh sand. 
Recovery of the gold was accomplished by 
precipitating with NaHS03 ; however, other 
methods of recovering the gold are avail- 
able. The byproduct metals cobalt and 
vanadium can be recovered from the iron- 
cobalt-vanadium solution using existing 
technology. 

TABLE 9. - Semiquantitative spectro- 
chemical analysis of gold-rhodium 
sand from reed switches 



Ele- 
ment 



Cone, 
wt pet 



Ag... 


. 0.003- 0.03 


Al... 


.003- .03 


Au. . , 


>10 


Co... 


.03 - .3 


Fe... 


.01 - .1 


Mg... 


.001- .01 



Ele- 
ment 

Mn.. 
Pb.. 
Rh.. 
Si.. 
Sn. . 



Cone, 
wt pet 

0.001- 0.01 
.003- .03 

>10 
.01 - .1 
.03 - .3 



51 



COPPER CEMENTATION WITH SELECTED MATERIALS FROM AV IONIC SCRAP (9) 



A process using either brittle aluminum 
base or magnetic metallics from elec- 
tronic scrap as a copper precipitant in 
acidulated CUSO4 solution was developed 
to separate and then upgrade a high-grade 
cement copper containing all or most of 
the precious metals. Aluminum-bearing 
obsolete avionic assemblies and plugs and 
connectors removed from electronic scrap 
were incinerated and then melted and cast 
to form brittle ingots. An assay of 
these incinerated and melted items is 
listed in table 10. The brittle ingots 
were broken up and rolls-crushed to minus 
35 mesh. Copper cementation took place 
when this material was agitated in the 
acidulated CUSO4 solution in a counter- 
current system. The precipitate con- 
tained better than 90 pet Cu and all of 
the associated precious metals. The ce- 
ment copper thus obtained can be bene- 
ficiated by fire refining to produce 
high-grade anode material for subsequent 
electrorefining to cathode copper. An 
anode mud is produced from which the pre- 
cious metals can be recovered using stan- 
dard procedures. 

TABLE 10. - Analyses of ingots produced 
by Incineration and melting of mili- 
tary electronic scrap 



Analysis, tr oz/st: 

Ag 

Au 

Concentration, wt pet 

Al 

Cu 

Fe 

Mn 

Ni 

Zn 

Insol 

^Not determined. 



The magnetic fraction of shredded elec- 
tronic scrap, which is mostly thin sili- 
con steel laminae from transformers, is 



Plugs and 
connec- 
tors 




Avionic 
assem- 
blies 



157.0 
0.16 

38.4 

24.0 

26.3 

(') 

0) 

0.23 

4.7 



an excellent copper cementation agent 
when it is agitated in an acidulated 
CUSO4 solution. It also contains nickel- 
alloy transistor caps that contain some 
gold and silver. A representative sample 
of the magnetic fraction from shredded 
electronic scrap was melted, cast, and 
sampled for assay. An assay of this ma- 
terial is listed in table 11. 

TABLE 11, - Analysis of a representa- 
tive sample of shredded magnetic 
scrap 

Value 

Analysis, tr oz/st: 

Ag 7,2 

Au 5,96 

Concentration, wt pet: 

Cr 0,7 

Cu 0,7 

Fe 75,8 

Ni 11,5 

Insol 4,4 

Copper cementation was most efficient 
when shredded magnetic scrap was used in 
a tumbler-type system, since a vigorous 
scrubbing action was needed to maintain 
the precipitation reaction. An analysis 
of the combined precipitate showed it to 
contain, in weight percent, 89 Cu, 1,1 
Fe, 0,32 Pb, 0,36 Sn, 1,5 insolubles, and 
in ounces per short ton, 4,0 Au and 11,5 
Ag. 

The total precipitate was melted in an 
induction furnace without flux. An an- 
ode, representing 90 pet of the charge 
weight, was obtained from the melt; the 
balance, containing most of the impuri- 
ties, remained as a sinter. The anode 
was placed in a polypropylene sock and 
electroref ined in a solution containing 
150 g/L H2SO4 and 40 g/L Cu at a current 
density of 12 A/f t^ , Previous electro- 
refining tests using these conditions 
produced a dense, smooth cathode deposit. 
Assays of the various products are listed 
in table 12, The anode mud can be fur- 
ther refined using standard procedures 
for gold and silver recovery. 



52 



TABLE 12. - Analyses of electro- 
refining products 





Anode 


Cathode 1 


Anode 
mud 


Analysis, 

Ag 

Au 


tr oz/st: 

• ••••••••• 


12.1 
4.3 

99.4 

0.13 

0.104 

0.047 

0.05 


ND 
ND 

99.9 

ND 
ND 
ND 
ND 


1,140 
412 


Analysis, 
Cu 


pet: 


57.0 


Fe 


ND 


Pb 

Sn 


0.7 
0.316 


Insol 


NS 



TABLE 13. - Analyses of products ob- 
tained from melting and electrolyti- 
cally solubilizing unreacted magnetic 
scrap residue 



ND Not determined. 

'a spectrographic analysis of the cath- 
ode copper showed the following impuri- 
ties (estimated): 0.001 wt pet Mg and 
0.01 wt pet Ca. 

The cleaned, unreacted residue from the 
copper cementation with shredded magnetic 
scrap (amounting to 11.0 pet of the orig- 
inal charge) was melted in an induction 
furnace and cast into an anode. An assay 
of the metal is listed in table 13. 

The anode was suspended in a elec- 
trolytic cell fitted with a graphite 



Analysis, tr oz/st: 

Ag 

Au 

Analysis, wt pet: 

Co 

Cr 

Cu 

Fe 

Ni 

Insol 



Resi- 
due^ 



35 
12 

0.7 

0.5 

11.0 

40.0 

46.4 

Na 



Anode 
mud 



239 
129 

NA 

NA 

55.0 

17.0 

18.0 

8.0 



Solution, 
g/L 



NA 
NA 

1.9 

0.9 

0.05 

50 

61 

NA 



NA Not analyzed. 

^Cleaned, unreacted scrap residue. 

cathode and partially solubilized in an 
electrolyte containing 150 g/L H2SO4. 
The metals in solution (table 13) could 
be treated by presently used technology 
to recover the nickel, cobalt, and chro- 
mium. The anode mud (table 13) can be 
treated using standard procedures for the 
recovery of the precious metals. 



SWEATING AVIONIC SCRAP TO PRODUCE ALUMINUM BULLION 
FOR FUSED SALT ELECTROLYSIS (10) 



The Bureau developed and tested two 
molten-salt electrorefining procedures 
for processing aluminxom ingots sweated 
from avionic scrap in order to recover a 
high-quality aluminum and concentrate the 
gold and silver in the aluminum-depleted 
anodes. An analysis of one of the ingots 
is listed in table 14. 

One system used a three-layer cell. 
The three molten layers were separated 
because of differences in the densities 
of the molten anode material, the molten 
salt electrolyte, and the molten refined 
aluminum. At the operating temperature 
range of 750° to 850° C, the approximate 
density was 3.3 g/cm-' for the molten 
electronic scrap, 2.7 g/cva? for the mol- 
ten electrolyte, and 2.3 g/cm^ for the 
molten refined aluminum. The electrolyte 
contained, in weight percent, 60 BaCl2, 
17 NaF, and 23 AIF3. The recovered alu- 
minum had a purity of 99.8 pet. The com- 
position of the anode residue from the 
three-layer cell (representing 33 pet of 



the electronic scrap ingot) is listed in 
table 15. The anode residue can be read- 
ily refined to copper bullion suitable 
for further treatment by aqueous acid 
electrolysis to separate precious metals 
and copper. 

TABLE 14. - Composition of electronic 
scrap ingot 

Value 

Analysis, tr oz/st: 

Ag 119 

Au 12 

Concentration, wt pet: 

Al 69.69 

Cu 19.98 

Fe 0.50 

Mg 0.13 

Mn 0.20 

Ni 0.22 

Pb 1.07 

Si 2.57 

Sn 1.19 

Zn 4.02 



53 



TABLE 15. - Composition of total anode 
residue, three-layer cell 

Value 

Analysis, tr oz/st: 

Ag 378 

Au 37 

Concentration, wt pet: 

Al 6.93 

Cu 61.61 

Fe 2.03 

Mg 0.007 

Mn 0.32 

Ni 0.61 

Pb 1.88 

Si 6.48 

Sn 6.75 

Zn 10. 16 

A second system used a compartmented 
cell that provided separate compartments 
for the anode and cathode metals instead 
of the density separation used in the 
three-layer cell. The cell was operated 
in the range of 750° to 800° C. The 
electrolyte consisted of an equimolar 
mixture of NaCl and KCl, with enough 
AICI3 to provide a 1- to 2-wt-pct Al 
concentration in the electrolyte. The 



recovered aluminum had a purity of 99.6 
pet. The composition of the anode resi- 
due from the compartmental cell (repre- 
senting 32 pet of the electronic scrap 
ingot) is listed in table 16. This anode 
residue can also be refined to copper 
bullion suitable for electroref ining. 
The two fused-salt electroref ining pro- 
cesses are technically feasible; however, 
economic evaluations have not been made. 

TABLE 16. - Composition of total anode 
residue, compartmental cell 

Value 

Analysis, tr oz/st: 

Ag 365 

Au 35 

Concentration, wt pet: 

Al 12.2 

Cu 60.9 

Fe 1.66 

Mg <0.01 

Mn 0.45 

Ni 0.68 

Pb 3.27 

Si 8.04 

Sn 3.64 

Zn 7.77 



INCINERATION, CAUSTIC LEACHING, SMELTING, AND 
ELECTROREFINING OF AVIONIC SCRAP (11) 



From a metallurgical standpoint, avi- 
onic scrap is a complex mixture of vari- 
ous metals, mostly copper, aluminum, and 
iron, attached to, covered with, or mixed 
with diverse types of plastics and ceram- 
ics. Precious metals occur as platings 
of various thicknesses, in relay contact 
points, on switch contacts and wires, and 
in solders. 

The plastics and organics can be elimi- 
nated prior to smelting by incinera- 
tion at 400° to 500° C in a gas-fired 
furnace. The furnace must be equipped 
with an afterburner and a scrubber in 
order to meet antipollution regulations. 
Iron alloys are removed with a drum 
magnet. The aluminum content, generally 
about 35 to 40 pet in avionics, may be 
removed by caustic leaching with regener- 
ation of the caustic. Leaching is not 
necessary unless there is a ready market 
for the AI2O3. Smelting of the inciner- 
ated scrap, with or without the aluminiim 



removed, is done to produce a homogeneous 
assayable product that can be sold to a 
custom smelter. An ingot produced from 
electronic scrap, without the aluminum 
removed, assayed 30 wt pet Cu, 18 wt pet 
Fe, and 32 wt pet Al, and 120 tr oz/st 
Ag and 1 tr oz/st Au. Smelting another 
batch of electronic scrap low in alumi- 
num produced an ingot that assayed 85 wt 
pet Cu, 4 wt pet Fe, and 0.2 wt pet Al, 
and 333 tr oz/st Ag and 26 tr oz/st Au. 
These two assays illustrate the hetero- 
geneity of electronic scrap composition. 
The remelted ingots, with or without the 
aluminum removed, can be further refined 
by slag additions and blowing air through 
the melt. The resulting copper bullion 
is refined electrolytically , and the an- 
ode slimes, which are rich in precious 
metals, are sent to a toll refiner for 
final processing. 

Economic evaluation of the various 
procedures for processing avionic scrap 



B lWjWIgWI IIHHg 



54 



indicates that incineration followed by 
smelting to form an assayable ingot 
and selling this product to a custom 



electroref iner is the most cost-effective 
procedure. 



TREATMENT OF SPENT TIN-LEAD SOLDER FROM MANUFACTURE OF ELECTRONIC 
PRINTED CIRCUIT CARDS TO RECOVER GOLD (12-13) 



Estimates indicate that 2,000 st of 
spent solder containing about 120,000 
tr oz Au is generated annually (12). 
One method for recycling the tin-led 
solder and recovering the gold is fused- 
salt electrolysis (12). Electroref ining 
of the spent solder is carried out at 
450° to 500° C with a molten KCl-SnCl2- 
PbCl2 electrolyte. The spent solder is 
charged to the anode container. Both 
anode and cathode electrodes are tung- 
sten. Electrolyte composition is, in 
weight percent, 14 KCl, 28 SnCl2 , and 58 
PbCl2, The electrolyte melting point is 
390° C, Refined tin-lead solder is re- 
covered at the cathode. In tests, gold 
at the anode increased from about 60 to 
2,200 tr oz/ St. The impure gold bullion 
collected at the anode is treated by con- 
ventional fire refining to recover rela- 
tively pure gold. 

A second method for recycling spent 
tin-lead solder is by drossing with alu- 
minum or zinc (13). The electronic sol- 
ders used in studying this method were 
nominal 60-40 tin-lead solders that had 
become so contaminated during wave sol- 
dering of printed circuit boards that 
they would no longer form acceptable 
bonds. The analyses of two contaminated 
solders used in the drossing study are 
listed in table 17. 

TABLE 17. - Analyses of some 
as-received scrap electronic 
solders, parts per million 




Recovery of gold from electronic sol- 
ders by phase separation requires a den- 
sity difference between the solder and 
the drossing agent. Scrap electronic 
solders have specific gravities ranging 
from 8.45 to 8.85 g/cm^ , while the spe- 
cific gravities of aluminum and zinc are 
2.70 and 7.14 g/cm^ , respectively. 

Samples of scrap solder were melted in 
open clay and clay-graphite crucibles in 
a conventional pot furnace. Bath temper- 
atures were held at 550° C for aluminum- 
treated solders and at 350° C for zinc- 
treated solders. Agitation of the bath 
during cooling helped phase separation. 
When the bath temperature reached 200° to 
250° C, the dross was removed from the 
solder by filtering. 

In addition to gold, the aluminum also 
removes significant amounts of antimony, 
copper, iron, and nickel from the solder, 
but is not effective for silver removal. 
Zinc is as effective as aluminum for re- 
moving gold from solder and also removes 
silver, copper, iron, and nickel, but not 
antimony. The use of zinc, however, im- 
poses distinct disadvantages. Compared 
with aluminum, approximately eight times 
as much zinc, on a molar basis, is needed 
to completely remove gold from the sol- 
der. Zinc dross does not become as vis- 
cous as aluminum dross as the bath tem- 
perature is lowered, thus complicating 
phase separation. In addition, solder 
dissolved larger quantities of zinc than 
aluminum, making it more difficult to re- 
purify the solder to an acceptable elec- 
tronic grade. Therefore, in general, 
aluminum drossing is preferred over zinc 
drossing except for solders containing 
substantial quantities of silver. 

Dross recovered by filtration contains 
essentially all of the gold originally 
present in the solder and typically rep- 
resents about 10 pet of the weight of the 
solder. This dross can be further pro- 
cessed to concentrate the gold, since 
about 85 pet of the dross is entrained 
solder. 



55 



Heating the dross in the range of 800° 
to 1,000° C will oxidize aluminum and 
aluminum-gold compounds together with 
some of the lead and tin. Most of the 
metallic solder entrained in the dross, 
amounting to 15 to 40 wt pet of the 
dross, remains in the metallic state and 
is separated from the oxidized material 
by grinding and screening. The metallic 
portion remains on a 200-mesh screen, 
whereas the oxidized material passes 
through. The entrapped solder recovered 
from the dross has approximately the same 
concentration of gold as the original 
scrap solder and is recycled when a fresh 
batch is treated. 

The gold in the oxide phase is in the 
free metallic state and can be reclaimed 
by aqua regia digestion or by a combined 
cyanidation-amalgamation procedure. If 
aqua regia digestion is chosen, the base 
metal oxides should first be treated to 
remove them so they do not report to 
the aqua regia with the gold. This is 
best accomplished by leaching the oxides 
successively with a 50-pct NaOH solution 
and concentrated HCl. The NaOH solution 
removes most of the AI2O3, while HCl 
dissolves the remaining soluble impuri- 
ties. After filtering and washing, the 
solutions are discarded. The remaining 



solids are treated in two steps with 
aqua regia, then filtered and washed. 
Approximately 98 pet of the gold dis- 
solves in the aqua regia. Gold can be 
recovered from the aqua regia solution by 
conventional cementation procedures. In 
cemented form, the gold can be easily 
processed by a gold smelter for final 
purification. 

The second technique utilizes combined 
cyanidation-amalgamation. Neither cyani- 
dation nor amalgamation alone is effec- 
tive because of the variation in the size 
distribution of the gold particles. The 
larger particles are removed by amalgama- 
tion, while the smaller particles are 
dissolved by the cyanide solution. In 
order to use the combined cyanidation- 
amalgamation procedure, the oxides must 
first be treated with 5N HNO3 . This 
treatment is necessary to remove a tar- 
nished surface film from the gold that 
hinders reaction of the gold with the cy- 
anide solution or with mercury. HNO3 
also dissolves unoxidized solder and some 
base metal oxides. Combined cyanidation- 
amalgamation is generally preferred over 
aqua regia digestion because of the high- 
ly corrosive nature of aqua regia, which 
complicates the selection of construction 
materials. 



CONCLUSIONS 



Efforts to recover precious metals from 
electronic and electrical scrap will 
benefit from initial hand segregation 
of the scrap material, in nearly every 
case, even if the segregation is only 
minimal. Base metals recovered will usu- 
ally pay for the segregation step. Me- 
chanical processing of electronic and 
electrical scrap by companies specializ- 
ing in this operations is available on a 
toll basis. The material fractions ob- 
tained from mechanical processing are 
homogeneous enough for assaying. Assays 



enable the generator or owner to deter- 
mine whether the scrap has enough value 
to at least pay for shipping and toll re- 
fining charges. The Bureau's studies in 
electronic and electrical scrap recovery 
have shown that mechanical processing to 
obtain homogeneous materials for assaying 
is beneficial and economical, and a study 
conducted by private industry has borne 
out this conclusion. However, further 
processing to recover precious metals is 
best handled by toll refiners. 



56 



REFERENCES 



1. Dunning, B. W. , Jr. Characteriza- 
tion of Scrap Electronic Equipment for 
Resource Recovery, Paper in Proceedings 
of the Sixth Mineral Waste Utilization 
Sjmiposium (IIT Res. Inst., Chicago, XL, 
May 2-3). ITT Res. Inst., 1980, 1978, 
pp. 403-410. 

2. Dunning, B. W. , Jr., and F. Am- 
brose. Characterization of Pre-1957 Avi- 
onic Scrap for Resource Recovery. Bu- 
Mines RI 8499, 1980, 20 pp. 

3. Ambrose, F., and B. W. Dunning, Jr. 
Precious Metals Recovery From Electronic 
Scrap. Proceedings of the Seventh Miner- 
al Waste Utilization Symposivim (ITT Res. 
Inst., Chicago, IL, Oct. 20-21, 1980). 
ITT Res. Inst., 1980, pp. 184-197. 

4. . Mechanical Processing of 

Electronic Scrap To Recover Precious- 
Metal-Bearing Concentrates. Ch. in Pre- 
cious Metals, ed. by R. 0. McGachie and 

Pergamon, 1980, pp. 67- 



A. G. 
76. 
5. 

H. V. 



Bradley. 



Dunning, B. W. Jr. , F. Ambrose, and 
Makar. Distribution and Analysis 
of Gold and Silver in Mechanically Pro- 
cessed Mixed Electronic Scrap. BuMines 
RI 8788, 1983, 17 pp. 

6. Hilliard, H. E. , B. W. Dunning, 
Jr. , and H. V. Makar. Hydrometallurgical 
Treatment of Electronic Scrap Concen- 
trates Containing Precious Metals. 
BuMines RI 8757, 1983, 15 pp. 

7. Hilliard, H. E., B. W. Dunning, 
Jr. , D. A. Kramer, and D. M. Soboroff . 



Hydrometallurgical Treatment of Electron- 
ic Scrap To Recover Gold and Silver. 
BuMines RI 8940, 1985, 20 pp. 

8. Hilliard, H. E., and B. W. Dun- 
ning, Jr. Recovery of Platinum-Group 
Metals and Gold From Electronic Scrap. 
Paper in Proceedings, 1983 International 
Precious Metals Institute International 
Seminar. The Platinum Group Metals — An 
In-Depth View of the Industry, ed. by 
D. E. Lundy and E. D. Zysk (Williamsburg, 
VA, Apr. 10-13, 1983). Int. Precious 
Met. Inst., 1983, pp. 129-142. 

9. Salisbury, H. B., L. J. Duchene, 
and J. H. Bilbrey, Jr. Recovery of Cop- 
per and Associated Precious Metals From 
Electronic Scrap. BuMines RI 8561, 1981, 
16 pp. 

10. Sullivan, T. A., R. L. deBeau- 
champ, and E. L. Singleton. Recovery of 
Aluminum, Base, and Precious Metals From 
Electronic Scrap. BuMines RI 7617, 1972, 
16 pp. 

11. Dannenberg, R. 0., J. M. Maurice, 
and G. M. Potter. Recovery of Precious 
Metals From Electronic Scrap. BuMines RI 
7683, 1972, 19 pp. 

12. Kleespies, E. K. , J. P. Bennetts, 
and T. A. Henrie. Gold Recovery From 
Scrap Electronic Solders by Fused-Salt 
Electrolysis. BuMines TPR 9, 1969, 8 pp. 

13. Ferrell, E. F. Recovering Gold 
From Scrap Electronic Solders by Dross- 
ing. BuMines RI 8169, 1976, 9 pp. 



T^U.S. GPO: 1985-605-017/20,138 



INT.-BU.O F MIN ES,PGH.,P A. 28195 



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